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Virginia Tech | Feed Distribution: The maximum yield from a spiral circuit can only be realized when the
same density cut point is maintained throughout the spiral circuitry. Unfortunately, one of the
common problems faced by coal spirals operated in industrial plants is poor feed distribution.
This problem can create differences in density cut points among different spiral units in a bank
and, in extreme cases, can lead to other operational issues such as beaching and sanding. Poor
feed distribution may also block the feed distributor port, which ultimately causes variations in
the slurry flow rates (Luttrell et al., 2003, 2007).
2.2.8 Spiral Flow Modeling
Although spiral concentrators are considered to be a mature technology, there is still
room for incremental improvements in design. The most likely approach for realizing these
design improvements is through the phenomenological modeling of the fluid flow patterns and
particle interactions that occur during spiral separations. Such fundamental modeling requires the
knowledge of operating variables (e.g., volumetric and dry solid feed rates), spiral design (e.g.,
pitch, diameter and trough shape) and particle properties (e.g., density, size and shape).
A number of researchers have developed spiral models based on mechanistic phenomena
(Holland-Batt, 1989; Atasoy and Spottiswood, 1995; Glass et al., 1999; Holland-Batt and
Holtham, 1992). These models have been used to predict the motion of particles in the flowing
film over the spiral trough surface. The first such model was developed by Holland-Batt (1989).
The model was capable of predicting the dynamics of fluid regimes and estimating particle
distributions across the spiral trough after separation. This model uses the Manning equation to
describe the primary flow in the inner region and the free vortex equation for the outer region
flow (Holland-Batt, 1989). This model is computationally intensive and its output is not directly
suitable for industrial process simulations (Li et al., 1995). The modeling work done by Holland-
33 |
Virginia Tech | Batt was further extended by researchers at JKMRC and their model was capable of predicting
operational performance for a given set of feed characteristics and splitter settings.
More recently, advances in computational fluid dynamics (CFD) and discrete element
modeling (DEM) of particle-particle interactions should ultimately help to completely investigate
the separation process and to design a better spiral geometry. During the last two decades, CFD
modeling in three dimensions has been used to simulate spiral flows (Holtham, 1990; Jancar et
al., 1995; Matthews et al., 1996). The concept of turbulence has also been incorporated into the
spiral modeling (Matthews et al., 1997). One of the most recent and most advanced models for
predicting particle partitioning during spiral separations was developed by Das et al. (2007). This
modified coal spiral model was based on three principal components, i.e., spiral geometry
modeling, fluid flow analysis and equilibrium force balancing for moving particles. This model
successfully predicted the radial equilibrium distribution of particles with respect to specific
gravity and particle size (Das et al., 2007). Nevertheless, while spiral models have provided
much insight related to the mechanics of particle separation, the ability of these fundamental
models to predict actual spiral performance for arbitrary applications remains a difficult task.
Consequently, there is a room for a more realistic analysis that would provide a truly quantitative
multiphase hydrodynamic description of spiral separations (Glass et al., 1999, Das et al., 2007).
Based on insight provided by fundamental modeling, Holland-Batt (1992) came up with
idea of rotating spirals. This research proposed that separation efficiency can be improved by
rotating the downward volumetric flow. It was argued that, in rotating spirals, one or more
additional forces were acting on the flowing film of particles, which results in a better separation
process. It was found that separation efficiency of fine feed particles increases from a spiral flow
34 |
Virginia Tech | that rotated over itself but, unfortunately, little or no improvement was found for the coarser feed
particles (Holland-Batt, 1992).
2.2.9 Spiral Circuitry
Until 1990, little work has been done in the area of optimizing spiral circuitry. This
situation changed rapidly in the 1990’s with the introduction of the compound spiral. The
compound spiral is essentially a two-stage middlings reclean circuit that operates along one
central spiral column (MacNamara et al., 1995, 1996), i.e., a short primary and short secondary
spiral are positioned on the same central tube. After the first stage, the primary reject is removed
through the central column and the primary clean and middlings are repulped and retreated on
the secondary spiral. Advantages of this design include lower density cut points, reduced floor
space, elimination of interstage pumping, and improved recovery (Weldon et al., 1997).
According to Bethell (2003), common spiral circuit configurations are as follows:
• Single stage spiral circuit without recycle.
• Spirals circuitry with middlings only recycles.
• Spiral circuit with clean coal only recycles.
• Spiral circuit configuration having both clean coal and middlings recycle.
Luttrell et al. (1998) used a linear circuit analysis technique to identify the optimum spiral
circuitry for compound spirals. They studied the following four different spiral circuit
configurations:
• Single stage spiral circuit.
• Conventional rougher-cleaner circuit without recycle.
• Modified rougher cleaner circuitry with middlings recycle.
• Rougher with middlings only recleaning.
35 |
Virginia Tech | Out of the above listed spiral circuits, the modified
rougher-cleaner circuit offered the best option for
improved performance, while maintaining a
reasonable circuit load. In the modified rougher-
cleaner spiral circuit, middlings particles from the
cleaner spiral are recycled back to the feed of
rougher spiral as shown by Figure 2.8. This
circuitry not only improves the separation
efficiency, but also reduces the density cut point
(Luttrell et al., 1998). Essentially all of the
compound spirals employed within the coal Figure 2.8 Modified rougher-cleaner
spiral circuit with middlings recycle
industry now use this two-stage configuation to (Luttrell et al., 1999). Used under fair
use, 2012.
reduce the entrappment of dense particles in the
high velocity flow region and to improve the sharpness of separation via the recycling of the
middlings product back to the feed stream.
2.2.10 Spirals for Ultrafine Coal Processing
A recent development related to the use of spirals is the introduction of ultrafine (0.15 x
0.044 mm) coal spiral circuits. In these circuits, deslimed ultrafine coal slurry is introduced to a
conventional spiral at a reduced dry solids feed rate of 0.45 to 0.50 t/hr/start and lower feed
percent solids of 10 to 12% by weight. These ultrafine coal spirals are reported to achieve a cut
point of 1.8 SG and an E value of 0.20. Moreover, the ash contents were reduced from 17.35%
p
to 9.84% and total sulfur from 3.56% to 3.0%. This spiral circuit was designed specifically to
treat oxidized ultrafine coal feed, which is difficult to upgrade by surface-based froth flotation
36 |
Virginia Tech | processes (Honaker et al., 2006). So far, the results suggest that this proprietary circuitry has the
potential to be effectively used to treat well liberated low ash ultrafine coal with high pyrite
contents, but may not suitable for poor feeds having high proportions of middlings or clay
particles (Luttrell et al., 2007).
2.2.11 Spirals for High Sulfur Coal Processing
Spirals are particularly well suited to reduce the sulfur contents of the final product in
coal applications (Zeilinger, 1976). Tavares and Sampaio (1990) reported that a standard LD-9
spiral reduced the clean coal sulfur content to 2.2% from 4.2%. While spirals commonly provide
a reject stream that is rich in sulfur, the absolute removal of sulfur from the clean coal product is
typically rather low in industrial operations. One obvious reason for this is that spirals can only
remove mineral, or pyritic, sulfur and not the organic sulfur associated with the carbonaceous
fraction of the coal. The poorer rejection may also be because spirals are mainly configured for
coal recovery rather than for pyrite removal (Kawarta et al., 2001).
On one hand hydrophobic coal pyrite particles (Oxidized pyrite particles are often
hydrophobic) or partially liberated pyrites often behaves in flotation like pure coal particles and
tend to report to the froth. On other hand spirals exploit the difference between specific gravities
of coal and rock forming minerals, thus the same hydrophobic pyrite particles can be rejected by
spirals because of their relatively high density. A two stage spiral in a rougher cleaner circuit
arrangement was found to improve the coal pyrite rejection by approximately 10% compared to a
single stage spiral (Kawarta et al., 2001). Tavares and Sampaio (1990) were showed that the
sharpness and density of separation increase with the increase in particle size.
37 |
Virginia Tech | 2.3 Teeter-Bed Separators
2.3.1 Introduction
Teeter-bed separators are hydraulic classifiers that have long been recognized as low-cost
and high capacity devices for both classification and density separation. A teeter-bed separator is
used for separating particles by size and/or density using a fluidized bed. A schematic diagram of
a typical teeter-bed separator is shown in Figure 2.9. Recently, many coal preparation plants in
the United States and in Australia have started to use teeter-bed technology as an alternative to
spiral separators for fine coal cleaning (Sarkar et al., 2008). Teeter-bed separators are based on
hindered settling and were originally used for particle size classification. However, if the feed
size distribution is within acceptable limits, these classifiers can be effectively used for the
concentration of particles based on differences in density (Bethell, 1988).
Teeter-bed separators have been manufactured since 1934. It was not until the early
1960’s, however, that this unit was for the first time used for coal recovery from waste piles and
tailings ponds (Drummond et al., 2002). Originally, teeter-bed separators primarily exploited
differences in coal particle size distributions for upgrading, but later developments in the
technology has now made it possible to separate particles primarily on the basis of density
differences. Recently, a new type of teeter-bed separators known as the reflux classifier has been
introduced to the coal and mineral industries. The reflux classifier technology has been reported
to achieve density cut points as low as 1.35 SG, while maintaining good separation efficiencies
that exceed those typically reported for coal spirals (Galvin et al., 2010).
38 |
Virginia Tech | 2.3.2 Particle Settling Theory
A solid particle falling under the influence of gravity in a viscous fluid is acted upon by
three forces, i.e., a downwards acting force of gravity, an upward acting buoyancy force due to
fluid displacement and an upward drag force acting in the direction of fluid flow. According to
Taggart (1945), the free settling of particles predominates if the percentage of solids by weight is
less than 15%. However, with an increase in the proportion of solid in the pulp, the settling rate
of solid particles began to decrease in response to an increase in the interstitial fluid velocity.
This condition, which is known as hindered settling, lower the settling rate compared to free
settling conditions. Littler (1986) states that hindered settling starts when the concentration of
solids in the pulp is about 20%. Under hindered settling conditions, a modified Newton’s law can
be used to determine the approximate falling rate of the particles (Wills and Napier-Munn,
2006). Mathematically, this condition can be represented by:
[2.1]
1/2
𝑣 = 𝑘[𝑑(𝐷𝑠−𝐷𝑝)]
where, is the falling velocity of particles, is the settling constant, d is the diameter of the
falling p𝑣article, D is the particle density D 𝑘a nd D is the pulp density. The hindered settling
S S P
phenomenon minimizes particle size classification and enhances density classification. The
hindered settling ratio between two different falling particles is always greater than the free
settling ratio.
With the increase in number of settling particles, a condition called “quicksand” is
reached. Under this condition, each particle is covered only by a thin layer of water and the
solids are in a state of “full teeter.” Every particle is free to move, but cannot do so without
colliding with other particles, so the particles tend to remain in place such that the mass of
39 |
Virginia Tech | discharged as underflow. Since the size and density of the feed particles are not uniform,
particles are generally segregated according to their mass within the teeter-bed (Luttrell et al.,
2006). The upward flow of fluidization water can be adjusted in such a way that the teeter-bed is
mainly composed of “near-density” particles suspended within the separator. The height of the
fluidized bed is normally controlled by on-line adjustment of the underflow rate using an
actuated valve. A transducer monitors the bed pressure and controls the valve opening and
underflow rate based on a specified set point value. Once steady-state conditions are reached,
small and light feed particles float upward along the rising current of water and report to the
overflow, while relatively coarse and denser particles sink and report to the under flow product.
Since their invention, teeter-bed separators have gone through significant advances in the
fundamental technology. As a result, there are many different classifying units, other than a
teeter-bed, that fall under this category of separator. These include Floatex density separators,
AllFlux separators, CrossFlow™ separators, HydroFloat™ separators and Reflux classifiers.
2.3.3 CrossFlow Separator
In order to maintain a high efficiency, non-turbulent conditions must exist within a teeter-
bed separator (Heiskanen, 1993). In a conventional teeter-bed, changes in flow patterns may
exist due to the introduction of feed into the teeter-bed chamber. This flow disruption can result
in the unwanted misplacement of particles. In order to overcome this problem, a modified teeter
bed known as the CrossFlow separator was designed the Eriez Manufacturing group. Figure 2.10
shows a conceptual diagram of a CrossFlow separator. As shown in this illustration, feed slurry
enters from the side of the CrossFlow unit and flows quiescently across the top of the teeter-bed.
This unique feed arrangement avoids turbulence normally associated with systems in which feed
slurry is injected below the pulp surface and into the center of the teeter-bed.
41 |
Virginia Tech | are too heavy to be carried by the rising teeter water and too light to penetrate the fluidized bed,
accumulate at the top of the teeter-bed. Eventually, more and more particles build up at the
surface of the bed, which forces some low-density coarse particles to enter the teeter bed.
Ultimately, these particles report to underflow along with the high-density particles. This
problem can be partially corrected by increasing the velocity of the fluidization water, but higher
velocities of can cause the misplacement of high-density fine particles. As a result, conventional
teeter-bed separators are inefficient when treating feed streams with a wide particle size
distribution and/or a narrow density distribution. To overcome this shortcoming, the HydroFloat
separator was developed by the Eriez Manufacturing group.
Figure 2.11 shows a conceptual diagram of HydroFloat separator. The HydroFloat is
essentially a teeter-bed separator into which small air bubbles are also introduced into the
fluidization water. If required, frothers and collectors may also be added to the teeter water aid in
bubble production and to render the desired mineral surfaces hydrophobic. The air bubbles attach
to hydrophobic particles within the teeter bed, which effectively reduce their density. Unlike
froth flotation, these bubble-particle aggregates have insufficient buoyancy to rise on their own,
but due to the attached bubbles, are light enough to rise to the top of the top of the teeter bed and
eventually report to the overflow launder. As a result, the HydroFloat combines the high capacity
of a teeter bed with the flexibility of a froth flotation process. The HydroFloat offers many added
advantages such as enhanced bubble particle contacting, better control of particle residence time,
lower cell turbulence and reduced air consumption (Mankosa, et al., 1999, 2003).
HydroFloat separators have been successfully used in phosphate and fine coal processing
operations (Barbee, 2007; Kohmuench et al., 2003; Mankosa et al., 2003). As shown in Figure
2.12, comparative testing of a HydroFloat separator on a deslimed coal (2 x 0.15 mm) provided
43 |
Virginia Tech | Figure 2.12 Test result obtained using 2 x 0.15 mm spiral feed from central Appalachia
(Mankosa et al., 1999). Used under fair use, 2012.
2.3.5 Reflux Classifiers
A reflux classifier is a teeter-bed separator that incorporates closely spaced parallel
inclined plates that accelerate particle movement (Figure 2.13). In terms of functionality, the
inclined plates are analogous to the inclined plates used in lamella thickener technology. For
density separations, the inclined plates help to suppress the effects of particle size and enhance
the separation of particles based on differences in relative density (Nguyentranlam and Galvin,
2004). A full-scale reflux classifier operated at flow rate of 16 tph of dry solids per square meter
of cross-sectional area treating fine (2 x 0.25 mm) coal achieved an overall density cut point of
about 1.7 SG and an E value to 0.15 (Galvin et al., 2004). Reflux classifier has also been used as
P
a dry fine coal cleaner in which air was used as the fluidizing medium.
45 |
Virginia Tech | possible to effectively separate small coal particles on the basis of density, with minimal size
effects, by using a vibrated reflux classifier having air sand dense medium. An overall E value
P
of 0.07 was achieved by dry processing coal (8 x 0 mm) in a semi-batch laboratory test using a
reflux classifier (Macpherson and Galvin, 2010; Macpherson et al., 2011).
2.4 Water-Only Cyclones
2.4.1 Water-Only Cyclone Description
A water-only cyclone (WOC) is a gravity-based separator that has been used in the coal
industry since the 1950's. The major advantage of a water-only cyclone over a dense medium
cyclone is that it does not require any external feed medium (Kim and Kalima, 1998; Patil et al.,
2007). On the other hand, water-only cyclone units are limited in terms of topsize to applications
involving only fine (<3 mm) coal feeds (Weyher and Lovell, 1969; Gottfried, 1978). A typical
water-only cyclone is a variant of hydrocyclone technology. The separator consists of a tapered
conical vessel that is open at its apex and joined from the top to a cylindrical section that
incorporates a tangential feed inlet (see Figure 2.14). The tangential inlet creates a rotating flow
within the unit that induces centrifugal forces that enhance the settling of larger and denser
particles to the cyclone wall and downward out the apex. A plate is fitted to the top of the
cylindrical section and an axially mounted overflow pipe passes through this plate. The overflow
pipe is extended into the body of the cyclone by a short removable section that is called a vortex
finder. The vortex finder prevents the short-circuiting of feed directly to the overflow. Smaller
and lighter particles are transported by the rotating flow to the center of the cyclone and up and
out through the vortex finder.
When feed is introduced tangentially under pressure to a water-only cyclone, the
movements of the particles inside the cyclone slow down because of the wide conical bottom.
47 |
Virginia Tech | This phenomenon creates a crowding of particles at the conical bottom. The crowded mass of
particles assists in developing hindered settling conditions and eventually in the formation of a
dense bed of particles. Thus, the separation process is based on the hindered settling velocity of
particles in a centrifugal field. The lighter particles are unable to penetrate the bed of higher
density material and thus report through the vortex finder and report as overflow, while denser
particles are discharged as underflow from the apex (Flintoff et al., 1987).
Figure 2.14 shows a schematic comparison between a classifying cyclone and a water-
only cyclone. Unlike conventional classifying cyclones, water-only cyclone units have a wide
conical angle and a long vortex finder that extends along the length of the cylindrical body.
Water-only cyclones utilize cone angles up to 120° and more, while classifying cyclones
designed for particle sizing commonly use cone angles of 10° to 20°. 2007. The larger cone
Figure 2.14 Comparison between classifying and water-only cyclone.
48 |
Virginia Tech | and on the size and specific gravity of feed particles. As shown in Figure 2.15, they proposed a
“generalized water distribution plot” that serves as a characteristic plot for the water split from
water-only cyclones. The slope of the plot represents the percentage of feed water reporting to
the overflow.
In a similar study, Majumder et al. (2011) studied the effects of various design and
operating variables on the separation efficiency of water-only cyclones used for fine (0.5 mm x
0) coal cleaning. This study revealed that vortex finder length, vortex finder diameter and cone
angle directly control the average residence time of coal particles inside the water-only cyclone.
It was also concluded that vortex finder diameter is the most sensitive variable among all the
variables studied.
In an earlier study, Kim and Klima (1998) suggested that for proper operation, the cone
angle of a water-only cyclone should be greater than 45°. However, too wide an angle (e.g.,
180°) reduces cyclone efficiency. At concentrations higher than 20% solids, the bypass of dense
particles to the overflow increases and, hence, reduces the separation performance. Cyclone
efficiency can also be enhanced through the use of multistage water-only cyclone circuitry.
Simulations performed by Kim and Klima (1998) indicated that a three-stage recirculating
overflow/underflow circuit can achieve ferrosilicon recoveries of 97% and quartz rejections of
95%.
The separation performance of a water-only cyclone is not only affected by its geometry,
but also depends upon the operating variables. Water-only cyclones are effectively used to
process coal streams finer than 1 mm, but coal feed up to 20 mm upper size have also been tried
on cyclones. Water only cyclones are most effective for a size fraction of 0.5 mm to 0.15 mm.
Particle size fractions finer than 0.15 mm will report to over flow irrespective of their density.
50 |
Virginia Tech | Bethell and Moorhead (2003) also studied the water-only cyclone/spiral circuitry to clean
fine (1 x 0 mm) coal. This study indicated that water-only cyclone and spiral circuits are flexible
and can be configured to operate at a high separating density to maximize carbon recovery or
may also be configured to operate at a low separating density to maximize coal quality. In order
to avoid excessive recycling rates that may occur in such circuitry, they proposed that the circuit
must either be configured with no recycle or with only a spiral middlings recycle. In the case of
treating finer feeds consisting of 0.6 mm top size, spiral middlings as well as clean coal can be
recycled without any fear of excessive recycling rates. However, recycling of both the spiral
clean coal and middlings must be restricted to circuitry where the water-only cyclone separating
density is sufficiently high to keep the recycle rate acceptable.
2.5 Froth Flotation
Conventional density-based separation processes are inefficient when used to upgrade
coal particles finer than about 100 mesh (0.15 mm). Thus, for this size range, froth flotation has
become the most commonly used coal processing technique. Flotation is a physico-chemical
separation process that utilizes differences in the surface “wettability” of coal and unwanted
rock/refuse particles. Figure 2.17 is a schematic diagram of a typical conventional flotation cell.
During operation, hydrophobic coal particles attach to air bubbles and are carried to the surface
and collected as concentrate, while hydrophilic particles remain in the aerated pulp and are
eventually discharged as tailings. A conventional, or mechanical, flotation cell consists of an
agitator and tank. The agitator keeps coal particles suspended and disperses the air bubbles
throughout the pulp. The agitation also provides turbulence within the pulp that promotes
collisions and attachment of hydrophobic particles to the air bubbles.
52 |
Virginia Tech | Flotation is a complex process that involves three phase flow and has been thoroughly
discussed by many authors (King, 1982; Schulze, 1984; Fuerstenau et al., 1985; Harris et al.,
2002; Rao, 2004). The process starts with the selective attachment of particles to air bubbles. The
bubble-particle aggregate rises into a froth phase due to buoyancy forces. Some of large particles
may be detached from the bubbles before reaching the froth. The entrainment of particles in the
water phase that reports to the froth may also occur. Unlike the selective process of flotation,
entrainment is not a selective and is detrimental to the grade of the froth product.
Column flotation is an important development in the froth flotation process. A column
cell consists of a long vertical cylinder into which air is added at the bottom. These cells are not
agitated by any mechanical means. Feed slurry, which is introduced at approximately two-thirds
of the way up along the column height, travels downwards through the column against rising air
bubbles. The selectivity in column flotation is enhanced via the use of water sprays that rinse
entrained mineral matter from the froth concentrate. Figure 2.18 is a schematic diagram that
compares the distribution of water in a conventional and column cell. Ideally, none of the water
from the feed slurry ever reports in the froth concentrate in column flotation. The column
technology was developed as an alternative approach to conventional flotation cells used in
cleaner circuits of mineral plants. Columns also have become increasingly popular for the
upgrading of ultrafine coal particles (Finch, 1995). There are now many types of column
flotation cells commercially available (Finch and Dobby, 1991; Jena et al., 2008). Two of the
technologies most commonly used in the coal industry include the Microcel column developed at
Virginia Tech and the CoalPro column developed by Canadian Process Technologies (CPT).
54 |
Virginia Tech | sulfur, diamond and talc naturally have high floatability and have high contact angles between
60° and 90° (Woods, 1994; Wills and Napier-Munn, 2006).
2.5.2 Flotation Reagents
It is theoretically possible to float non-polar hydrophobic minerals such as coal without
the aid of chemicals. However, the flotation process generally requires reagents be added to
promote bubble formation and enhance kinetics. These reagents can be classified as collectors,
frothers and modifiers (or regulators). Collectors are organic compounds that create or enhance
the hydrophobicity of selected particles so as to facilitate bubble-particle attachment. Frothers
are added to stabilize the formation of bubbles in the mineral pulp. Frothers also help to maintain
a stable froth that increases flotation kinetics and allows for the selective drainage of entrained
gangue minerals (Barbian, 2005; Melo and Laskowski, 2005). Modifiers, or regulators, are often
used in the flotation industry to control or modify the action of collectors. These reagents can be
generally classified as activators, depressants and pH modifiers. Activators are typically
inorganic substances that alter the chemical nature of the selected mineral surfaces and help to
make these particles hydrophobic due to the action of collectors (Zhou and Chander, 1991).
Depressants are used to increase the selectivity of the flotation process. Depressants prevent the
flotation on unwanted minerals by suppressing hydrophobicity. The selective separation that
occurs during flotation is also dependent on a delicate balance between the concentrations of the
reagents and the pulp pH (Buckley and Woods 1997; Raston et al., 2001; Wills and Napier-
Munn, 2006). Most industrial coal flotation systems operate over pH values between 6.5 and 8.
56 |
Virginia Tech | 2.5.3 Release Analysis
Flotation is not a density-based separation process, thus traditional washibility analysis
involving the density partitioning of particles cannot be directly applied to the flotation process.
In order to characterize the ideal flotation separation process, a method called “release analysis”
is frequently used by the industry. This technique was developed by Dell (1953) as the
equivalent in froth flotation to float-sink analysis in gravity concentration. In Dell’s release
analysis procedure, all the floatables particles are initially separated from non-floatable particles
by repetitive recleaning of the froth product (see Figure 2.19). After removing all of the non-
floatable particles, the remaining froth concentrate is repulped and then sequentially recovered in
small increments under conditions of steadily increasing flotation intensity. In most cases, the
flotation intensity is usually controlled by adjusting aeration rate and impeller speed. This
procedure typically produces a series of incremental froth products that range from highest purity
(recovered in the first increment) to lowest purity (recovered in the last increment).
Figure 2.19 A typical release analysis test.
57 |
Virginia Tech | 2.5.4 Coal Flotation
Froth flotation is generally used in coal preparation for upgrading the finest fractions of
coal in size ranges below about 100 mesh (0.15 mm). Coal particles are naturally hydrophobic
and require little addition of collector, although petrochemical products such as kerosene or
diesel oil are often added to enhance the hydrophobicity. Frother dosages are usually high to
keep the froth mobile (Wills and Napier-Munn, 2006). Coal floatability can be affected by
maceral type, mineral content and surface oxidation. Problems associated with the flotation of
oxidized coal include poor recovery, high ash contents of clean coal and higher reagent doses.
Moreover, mixing oxidized coal with a good coal may hurt the floatability of the good coal as
well (Luttrell, 2011). Pyrite laden coal also adds more complexities to the flotation process
(Kawarta, 2001).
Although most modern U.S. plants only treat coal feeds finer than about 0.15 mm, some
installations do exist in which particle sizes as large as 0.6 mm are effectively upgraded by this
process. Unfortunately, the exact relationship between coal particle size and the flotation rate is
difficult to understand. In most cases, the flotation rate initially increases with particle size,
reaches a maximum plateau value, and then decreases afterwards with a further increase in the
coal particle size (Taweel, 1986). The poor floatability of coarser coal particles may be due to
higher detachments rates resulting from the increased mass of larger particles (Lynch et al.,
1981). Mineral particles attached to the coal, which are commonly referred to as slime coatings,
can also decrease the floatability of the associated coal particles.
Coal flotation circuits are relatively simple and typically only require roughing and
scavenging circuits, although roughing circuits are usually sufficient in most industrial
operations. Generally, two types of fine flotation circuits are currently in use by the U.S. coal
58 |
Virginia Tech | preparation industry: conventional 0.15 mm x 0 circuits and deslime 0.15 x 0.045 mm circuits.
Typically, classifying cyclones are also incorporated into the flotation circuit to make a cut at
0.15 mm for conventional circuits or at 0.15 mm and 0.045 mm for deslimed circuits. In deslime
circuits; the minus 0.045 mm material is discarded waste. It is because increased amounts of coal
fines overloaded the carrying capacity of flotation cells and cause the loss of coarser particles
form the bubbles due to preferential loading of the finer sizes. Phillips I., Dennis. (1998). To
produce higher quality froth products, froth washing is often necessary and this can also be
achieved using Jameson cells and flotation columns (Nicol, 2000). Although the design and
operation of column flotation cells is more complex as compared with that of conventional
flotation cells, the steady decline in high-grade feedstocks will likely force operators to utilize
this technology in order to meet increasingly stringent customer demands (Yoon et al., 1997;
Kohmuench et al., 2007).
59 |
Virginia Tech | CHAPTER 3 PERFORMANCE COMPARISON OF FINE COAL
CLEANING ALTERNATIVES
3.1 Abstract
While dense medium processes have largely become the standard approach for treating
coarse coal, the types of unit operations used to upgrade fine (<1 mm) coal continues to vary
substantially from plant to plant and across different geographic regions. In light of this disparity,
an experimental study was undertaken to compare the separation performance of some of the unit
operations commonly used to upgrade fine coal streams. Processes examined in this pilot-scale
study included spirals, water-only cyclones, teeter-bed separators and froth flotation. Each
cleaning technology was tested on the coal feed from the same source. To fairly compare each
process, size-by-size separation efficiencies were determined for each process from characteristic
recovery/rejection curves. The resulting data was used to identify size ranges most appropriate
for the various alternative processes.
3.2 Introduction
Modern coal preparation facilities incorporate a wide variety of solid-solid separation
processes for coal upgrading. Dense medium processes, which include dense medium vessels
and dense medium cyclones, have become the preferred method for treating coarse coal in most
new plants (Figure 3.1). The widespread acceptance of dense medium technology can be
attributed to its large capacity, high efficiency and operational flexibility. In contrast, a variety
of commercially viable flowsheet configurations exist for treating coal finer than 1 mm (Honaker
et al., 2007). These circuit configurations may include various combinations of water-based
density separators such as spirals and water-only cyclones as well as various types of surface-
72 |
Virginia Tech | 3.3 Separation Efficiency
One of the difficulties associated with the comparison of fine coal cleaning processes is
the selection of a suitable performance indicator. For example, experimental test data often
consist of paired groups of values such as clean coal yield and clean coal ash. A cleaning process
that gives both a higher yield and a lower ash is obviously the best choice. On the other hand, the
choice of which process is superior is not so obvious when one unit provides a higher yield while
the other gives a lower ash. Ideally, these paired data sets need to be reduced down to a single
numerical performance indicator that can simultaneously indicate how effectively carbonaceous
matter is recovered versus how efficiency the ash (or other quality indicator) was rejected from
the feed. An indicator such as organic efficiency is well suited to this purpose for density-based
separation processes, but requires extensive float-sink analysis that may be cost prohibitive and
inappropriate for very fine particles unless special methods are employed (i.e., centrifugal float-
sink testing). Likewise, an arbitrary performance measure, such as the clean coal yield minus the
clean coal ash, is also undesirable since it has no real physical meaning, no fixed upper or lower
limits, and no definable relationship with an ideal separation.
The problem of assessing the performance of a separator using a single indicator has been
addressed in the literature (Stevens and Collins, 1961; Schultz, 1970; Salama, 2001). According
to these studies, the separation efficiency for a process should be defined as the theoretical
percentage of feed material that passes through an ideal separation. Mathematically, the
separation efficiency can be calculated as the recovery of desirable material in a given product
minus the recovery of undesirable material in the same product. For the case of coal, the
separation efficiency (E) can be obtained from:
75 |
Virginia Tech | [3.1]
E = R−(100−J) = R+J−100
in which R is the combustible recovery and J is the ash rejection. R represents the percentage of
combustible matter present in the feed that reports to clean coal, while J represents the
percentage of ash present in the feed that reports to reject. These two important performance
indicators can be calculated from the ash contents of the feed (f), clean coal (c) and refuse (r)
streams using:
[3.2]
(f−r)(100−c)
𝑅 = 100(c−r)(100−f)
[3.3]
(f−c)r
J = 100(r−c)f
Combining Eqs. [3.1]-[3.3] yields:
[3.4]
r−f 100−c c
E = 100r−c�100−f − f�
The leftmost term (i.e., 100[r-f]/[r-c]) in this expression is the total mass yield of clean coal,
while the terms (100-c)/(100-f) and (c/f) represent the concentration ratios of combustibles and
ash, respectively. Therefore, for the case of coal, the separation efficiency is comprised of the
clean coal yield times the difference in the concentration ratio between combustibles and ash.
The important relationship defined by Eq. [3.4] can be plotted graphically as shown in
Figure 3.2. A simple splitter, which gives no selective separation, is represented in this plot by a
diagonal line passing between the top-left corner (i.e., 100% recovery and 0% ash rejection) and
bottom-right corner (i.e., 0% recovery and 100% ash rejection). In contrast, a perfect separation
is represented by the single point in the top-right corner of the plot (i.e., 100% combustible
recovery and 100% ash rejection). These two boundaries represent separation efficiencies of 0%
76 |
Virginia Tech | [3.5]
∗ 100
R 100
c = �100−J�� f∗ −1�+1
[3.6]
∗ 100
100−R 100
r = � J �� f∗ −1�+1
This ability is particularly useful when comparing different sets of experimental data for which
the feed ash has changed slightly. Without this normalization step, it is difficult to distinguish
whether a superior yield-ash point is due to a true enhancement in separation performance or just
an artifact of a lower feed ash content.
3.4 Experimental
A pilot-scale test circuit was constructed for the purpose of evaluating several different
fine coal cleaning processes. Unit operations examined in the experimental program included a
two-stage compound spiral, teeter-bed separator, HydroFloat separator, water-only cyclone and
froth flotation cell. Specifics related to the working features of the teeter-bed separator and
HydroFloat separators have been discussed elsewhere in the literature (Kohmuench et al., 2001,
2002). A simplified schematic of the closed-loop test circuit is shown in Figure 3.3.
During testing, a coal slurry mixture was prepared by adding water and dry coal of the
desired particle size into a 2-m diameter feed sump. For most tests, the feed coal consisted of
nominal 1 x 0.15 mm solids, although coarser splits containing either 2.3 x 1 mm or 2.3 x 0.15
mm were also used in selected tests. The particle mixture was held in suspension using a 25-cm
diameter blade mixer. Slurry from the sump was pumped at a controlled rate using a variable-
speed centrifugal pump equipped with a 35-cm diameter impeller. If necessary to maintain
adequate particle suspension and slurry mixing, some portion of the slurry was returned back to
the sump via a bypass valve. Feed slurry from the pump was passed up to an upper level floor to
each unit operation being testing.
79 |
Virginia Tech | collected by diverting the full flow of the product streams into sample containers. For most of the
test units, the collected products included timed samples of clean coal and refuse. However, the
two-stage compound spiral included a product box partitioned to collect six different samples
across the profile of the second stage spiral as well as an upper draw-off point for the collection
of primary refuse from the first stage spiral. This configuration made it possible to
simultaneously collect timed samples of clean coal, refuse and five different middlings products
so that complete recovery-rejection curves could be generated for each spiral test run. After
collecting and weighing the slurry samples, the solids were filtered, dried, weighted and analyzed
for ash content. The experimental data were then evaluated using a spreadsheet-based mass
balance routine to ensure that reliable data had indeed been obtained for each test run (Luttrell,
2004).
3.5 Experimental Results
3.5.1 Spiral Testing
Spiral testing was conducted using a two-stage compound spiral (Multotec SX7). The
unit was operated in accordance with recommended guidelines reported in the literature (Luttrell
et al., 2003; 2007). As indicated previously, the 1-m diameter commercial-scale unit was
equipped with a partitioned collection box so that seven products could be simultaneously
collected across the spiral profile. Figure 3.4 shows the size-by-size recovery-rejection curves
obtained using the spiral under standard operating conditions of 8.6 m3/hr (38 GPM) and 2.5 t/hr
(2.8 TPH). The data is re-plotted in Figure 3.5 to better illustrate the cumulative effect of splitter
position across the second stage spiral profile on separation efficiency. Position “P1” represents
the cleanest product taken at the outer most position across the spiral profile, while “P6”
represents the high ash reject product taken at the inner most position near the center support
81 |
Virginia Tech | (200 mesh). These results were not unexpected since spirals are typically utilized to upgrade
coal feeds in the 1 x 0.15 mm size range. It is also interesting to note that the separation
efficiency for the 0.3 x 0.15 mm size class is comparatively low for products collected from
splitter positions “P2” and lower, but improves substantially as middlings products “P3” and
“P4” are added to the combined clean coal product (Figure 3.5). The high separation efficiency is
largely driven by the very high combustible recovery obtained for this size fraction when
collecting products “P1” through “P4”.
Figure 3.6 shows the size-by-size effect on separation efficiency of reducing the spiral
volumetric feed flow rate to only 5.7 m3/hr (25 GPM). When compared to the plot (Figure 3.5)
for the higher flow rate of 8.6 m3/hr (38 GPM), the reduction in flow resulted in substantial
decreases in the separation efficiency of solids in the plus 1 and 1 x 0.6 mm size classes
reporting to the middlings and reject streams (“P3” through “P6”). Close examination of the
experimental data indicated that the efficiency reduction was due to the loss of coarser coal
particles resulting from that shift of volumetric flow to a lower point in the spiral trough. In other
words, the lower flow rate reduced the density cutpoint for the coarser solids to a value lower
value. A similar large decline in separation efficiency was not observed for coarser solids in the
clean products represented by products “P1” and “P2”, although a small reduction in separation
efficiency did occur. Only slight changes in separation efficiency were observed for solids of
intermediate size ranges of 0.6 x 0.3 mm and 0.3 x 0.15 mm. On the other hand, the lower flow
rate resulted in significant increases in separation efficiency for finer solids in the 0.15 x 0.074
mm and 0.074 x 0.044 mm size fractions. As such, the data suggests that lower flow rates are
preferred for the separation of finer solids (<0.15 mm), while larger flow rates are preferred for
the upgrading of coarser solids (>0.3 mm). However, since most coal feeds to spirals are
84 |
Virginia Tech | 3.5.2 Water-Only Cyclone Testing
Figure 3.7 and Figure 3.8 show the size-by-size separation performance of the water-only
cyclone evaluated in the pilot-scale tests. In this particular case, the water-only cyclone was
operated at three different feed solids contents ranging from a low of 8% to a high of 15%. In
general, poorer separation efficiencies were observed for the water-only cyclone across all size
fractions when compared to the spiral separation curves. One explanation for the lower
separation efficiencies was that the geometry of the water-only cyclone was not ideally
optimized for the type of feed coal used in the pilot-scale test program. On the other hand, the
experimental data from this test program and others reported in the literature do indicate that
high separation efficiencies are more difficult to maintain using this technology. The data show
large variations in separation efficiency across each size fraction, suggesting that the density
cutpoint for each size fraction declines sharply as particle size increases. This notable difference
in performance is probably a major contributing factor in the historical shift in operator
preferences from water-only cyclones to spirals over the last several decades in the U.S. coal
preparation industry. Still, acceptable separation efficiencies in the range of 35-40% were
achieved for the coarsest size fractions (>0.6 mm) when operating at the water-only cyclone at
the highest feed solids content of 15%. It should be noted, however, that the efficiency declined
sharply as the feed solids content dropped from 15% solids to 12% and further to 8% solids. This
finding was not unexpected since the reduction in feed solids content shifted the operation of the
cyclone from that of a density-based separator to that of a particle size separator (i.e., classifying
cyclone). As such, the data collected under the current test program suggest that water-only
cyclones should not be operated as a standalone process due to the inability of this process to
simultaneously maintain good separation efficiencies across all size fractions. In addition, to
86 |
Virginia Tech | interstitial liquid velocities that resist the penetration of the slow settling particles. As a result,
small/light particles accumulate in the upper section of the separator and are eventually carried
over the top of the device into a collection launder. Large/heavy particles, which settle at a rate
faster than the upward current of rising water, eventually pass through the fluidized bed and are
discharged out one or more restricted ports through the bottom of the separator. The HydroFloat
is a special type of teeter-bed separator in which small air bubbles are also injected to avoid the
loss of larger high-mass particles to the underflow, which is not uncommon for teeter-bed
separators. In the current test program, the teeter-bed units were tested at three different
elutriation water rates. Only the best set of test data is shown for the standard teeter-bed unit.
Each unit was configured to run under operational conditions as recommended by the equipment
manufacturer.
The data plotted in Figure 3.10 shows that the standard teeter-bed separator was capable
of providing high separation efficiencies above 60% for size classes larger than 0.6 mm. In fact,
the solids contained in the 1.7 x 0.6 mm size fraction where cleaned at a separation efficiency of
about 74%, which exceeded the separation efficiencies obtained using the spiral technology. The
separation efficiencies were generally further improved when using the HydroFloat technology.
In particular, the separation efficiency of the coarsest material in the plus 1.7 mm fraction
increased from about 64% to nearly 82% using the HydroFloat separator, while the separation
efficiency for the 1.7 x 1.0 mm fraction increased from 74% to just over 80%. The data suggest
that the injection of air into the teeter-bed reduced the likelihood that coarser high-mass particles
would report to the underflow stream and be rejected, i.e., it increased the density cutpoint for
the coarser size fractions. Relatively little difference in separation efficiency was noted between
the two technologies for particles contained in the 1 x 0.6 mm size fraction. Surprisingly, the
89 |
Virginia Tech | minute or greater, the test data surprisingly indicate that very good separation efficiencies of near
70% could be obtained for particles in the coarse size ranges of 1 x 0.6 mm and 0.6 x 0.3 mm.
For all other size classes, somewhat lower separation efficiencies of 60-64% were obtained.
Unfortunately, practical experience suggests that the high separation efficiencies obtained in the
laboratory tests for particles larger than 0.6 mm would probably not be possible in a full-scale
industrial application (Moxon et al., 1988; Laskowski et al., 2007). Larger cells used in industrial
applications are typically unable to recover larger particles due to froth transport problems. On
the other hand, experience indicates that laboratory performance data obtained for particles finer
than 0.3 mm can typically be duplicated in industrial scale machines. Therefore, the data
obtained for particles larger than 0.3 mm from the laboratory tests should be substantially
discounted (perhaps by 30-50%) when compared to the other data reported in this study.
92 |
Virginia Tech | unit. The feed size distribution evaluated in each unit ranged from a top size of 2.3 mm down to
a bottom size of zero. It should also be noted that two sets of spiral data are included in the
comparison plot for feed top sizes of 2.3 and 1.0 mm, respectively. This step was necessary since
the removal of 2.3 x 1.0 mm solids from the spiral feed was found to provide a significantly
improved separation efficiency curve that needed to be considered in the process comparison.
Several interesting observations can be made from the comparative performance data. For
example, the data show that several different processes appear to be capable of providing good
separation efficiencies for particles larger than 0.6 mm. These units include spirals, teeter-bed
and HydroFloat separators. Of these, the HydroFloat tended to be the most robust at maintaining
the separation efficiency as the particle size increased into the plus 1.7 mm size range. The
injection and attachment of small air bubbles to the coarser coal particles avoided the decline in
recovery and corresponding reduction in separation efficiency that occurred in the standard
teeter-bed separator operated without air injection. Unfortunately, the performance of the teeter-
bed units (standard and HydroFloat) dropped sharply from about 70-75% down to 45-55% for
the 0.6 x 0.3 mm size class and down to unacceptably low values of less than 20% for the 0.3 x
0.15 mm size class. Particles finer than 0.3 mm are simply too small to overcome the interstitial
velocity of fluid in the teeter-bed and report to overflow regardless of quality. In contrast, the
spiral tended to maintain a reasonably good separation efficiency in the range of 55-60% for the
size class as small as 0.3 x 0.15 mm. As such, spirals appear to be a very good choice for treating
feeds with a large proportion of solids in the 1 x 0.6 mm and 0.6 x 0.3 mm size ranges.
As should be expected, the experimental results also demonstrated that froth flotation
should be the preferred technology for upgrading particles finer than 0.3 mm and the only
realistically viable process for upgrading particles finer than 0.15 mm. Above a critical particle
95 |
Virginia Tech | size of about 0.3 mm, flotation performance diminishes due to the inability of commercial
machines to effectively recovery larger particles as a result of issues associated with pulp-froth
transport. Interestingly, spiral separators appear to be almost as good as flotation for the 0.3 x
0.15 mm size fraction, so the decision as to whether to treat this size class by flotation or by
spirals would partially depend on site specific considerations such as the inherent floatability
characteristics of the feed coal. For example, flotation may be the most attractive approach for
treating the 0.3 x 0.15 mm fraction for a high-rank easily floated coal, while spirals may be a
much better choice if the feed material in this size fraction responds poorly to flotation due to
poor floatability as a result of factors such as weathering and surface oxidation. With the
exception of spirals, none of the other density-based processes examined in this study provided
separation efficiencies higher than about 15-20% for particles smaller than 0.3 mm. As such,
these processes are not recommended for this size range or for feeds containing a large portion of
solids in this size range.
96 |
Virginia Tech | CHAPTER 4 ENGINEERING DEVELOPMENT OF THE EXPANDED
STAGE COMPOUND SPIRAL CIRCUIT
4.1 Abstract
An in-plant experimental study was performed to evaluate the separation performance of
five different spiral circuit configurations. The spiral circuit feed consisted of nominal 1 x 0.15
mm particles from the underflow of a bank of classifying cyclones. The experimental data
obtained from the in-plant study was mass balanced using spreadsheet-based routines, evaluated
and compared for separation efficiency, clean coal yield, organic efficiency and combustible
recovery. On the basis of this investigation, it was determined that the best performance could be
achieved using a new four-stage circuit in which the clean coal passes through all four stages,
high-ash refuse is removed after each of the four stages, and middlings from the second and final
stages are recycled back to the original feed. This circuit provided the best separation efficiency,
cumulative clean coal yield and combustible recovery among all the other spiral circuits tested.
At the same clean coal ash, the new spiral circuit increased the cumulative clean coal yield by
1.9 % as compared to that achieved using the existing two-stage compound spiral currently
installed at the plant. The experimental work also proved that the repluping after two turns of
spirals is not effective in improving separation performance.
4.2 Introduction
The past two decades has witnessed widespread use of spirals to clean fine (1 x 0.15 mm)
coal. Modern coal preparation plants incorporate spiral separators in a variety of different circuit
configurations. The literature contains numerous studies that have examined the effects of spiral
design variables (such as spiral construction, spiral pitch and diameter, and spiral length) and
101 |
Virginia Tech | spiral operating variables (such as feed percent solids, dry feed rate and volumetric flow rate) on
the spiral separation efficiency. Unfortunately, comparatively little research has been conducted
to optimize spiral circuitry in relation to separation efficiency. Perhaps the first studies on coal
spiral circuitry were started in 1992 when Holland-Batt came up with the idea of rotating spirals.
This research proposed that separation efficiency could be improved by rotating the downward
volumetric flow on the spiral trough. In rotating spirals, one or more additional forces were
acting on the flowing film of particles, which results in a better separation process (Holland-Batt,
1992). It was found that separation efficiency of fine feed particles increases from a spiral flow
that rotated over itself but, unfortunately, little or no improvement was found for the coarser feed
particles (Kohmuench, 2000).
During the late 1990’s, researchers at Virginia Tech utilized the linear circuit analysis
technique to improve the separation efficiency of a spiral circuit. They concluded that a reduced
gravity cut point and improved separation efficiency could be achieved with a rougher-cleaner
spiral circuitry (Figure 4.1), provided that the middlings were recycled back to the spiral circuit
feed (Luttrell et al., 1998). The invention of the compound spiral was another important
milestone in the coal spiral circuitry. The compound spiral is essentially a two-stage spiral where
a short primary spiral and a short secondary spiral are mounted around the same central column.
After initial separation on primary spiral, reject is removed through a primary refuse cutter and
the remaining slurry is captured, remixed and reintroduced on the following secondary spiral.
Finally, after a recleaning stage on secondary spiral, the products are collected as clean coal,
middlings and secondary refuse by diverting the flow through appropriate product splitters. Both
primary and secondary refuse are discarded, while the middlings are discarded, added to the
cleaned coal product, or recycled back with the original spiral feed, depending on the spiral
102 |
Virginia Tech | Figure 4.3 provides actual separation efficiency curves for one of the experiments
conducted using the compound spiral test rig. The data shows that the separation efficiency at
splitter position “SC1” is substantially lower than the separation efficiency for the same size
class at splitter position “SC2”. A dashed oval shape is drawn in Figure 4.3 to show this
unexpected decrease in the separation efficiency at position “SC1”. In other words, the clean coal
product collected at position “SC1” contained more misplaced rock particles than the product
collected at position “SC2”. This work verified that some unwanted high-ash particles of rock
are trapped in the outer high velocity flow region of a spiral separator.
To improve the separation performance of a spiral circuit, five different full-scale spiral
units were installed, experimentally tested and compared in an industrial coal preparation plant.
The separation performance of each spiral circuit was evaluated by comparing size-by-size
separation efficiencies, cumulative clean coal yields, organic efficiencies and combustible
recoveries. This article describes the layout of all five spiral circuits, provides details related to
the experimental test program, and summarizes the test results obtained from the comparative
evaluation of the separation performance of all the experimentally tested spiral circuits.
105 |
Virginia Tech | Figure 4.4 Cardinal Preparation Plant, WV.
online ash analyzer. The plant incorporates three identical 700 TPH capacity modules. The coal
feed consists of high-quality bituminous coal mined from nearby areas. A double-deck banana
screen sizes the run-of-mine feed coal at 12 mm on the top deck and then at 1 mm on the bottom
deck. The coarse feed goes to dense medium vessels and dense medium cyclones for density
separation. The fine (-1 mm) coal processing circuit consists of classifying cyclones, spirals,
flotation columns and screenbowl centrifuges. The spiral circuit, which was the focus of the
current investigation, is fed with fine (1 x 0.15 mm) coal from a bank of 15-inch diameter
classifying cyclones. The cyclone underflow passes into a distributor that simultaneously feeds
six sets of triple-start compound spirals. The middlings from the spiral circuit is recycled back to
the spiral feed (Figure 4.5). Column flotation is used to process deslimed (0.15 x 0.044 mm) feed
coal, while the ultrafines (0.044 mm x 0) is discarded as waste. Finally, the combined deslimed
spiral and column flotation concentrates are mixed together and dewatered using several 42 x
144-inch screenbowl centrifuges (Bethell and DeHart, 2006).
107 |
Virginia Tech | leaving the spiral circuit. A numbered subscript represents an internal stream and it also indicates
the spiral number where the product stream (clean, tailings middlings) reports. The use of the
value “0” in the subscripts indicates that no such stream exists. For example, consider the simple
notation of “1 ”. In this case, the number “1” means that circuit consists of a single spiral unit.
CT0
The subscripts C, T, and 0 represent the clean, tailings and middlings streams, respectively.
Moreover, these subscripts also indicate that two products (clean and tailings) leave the spiral
circuit and there is no middlings stream produced by the circuit.
Using this notation system, a standard compound spiral circuit (without middlings recycle
back to the original feed) is represented by “1 + 2 ”. The information that can be inferred
2T0 CTM
from this notation is as follows.
• The spiral circuit consists of two spiral units.
• The designation “1 ” represents the first-stage spiral and “2 ” represents the second-
2T0 CTM
stage spiral.
• The number “2” associated with the first-stage spiral indicates that the clean stream
produced by this spiral is an internal stream that reports to the second-stage spiral.
• The letter “T” associated with the first-stage spiral indicates that the tailings leave the
spiral circuit as an external product stream.
• The number “0” indicates that no middlings stream is produced by the first-stage spiral.
• The subscripts C, T and M associated with the second-stage spiral indicates that three
product streams are produced by this spiral and leave the circuit as clean, tailings and
middlings, respectively.
110 |
Virginia Tech | 4.3.4 Experimental Circuit Configurations
An extensive experimental testing program was carried out on five different spiral circuit
configurations using full-scale spirals. All the experiments were performed using standard spiral
operating conditions, i.e., a volumetric slurry flow rate of 38-40 GPM and a dry solids feed rate
of about 2.5-2.7 TPH per start. Table 4.1 shows the operating conditions for all experiments.
Table 4.1 Summary of the design and operating parameters.
Design and Operating Parameters for all Spiral Circuits
Spiral
Spiral Number Number of Feed Feed Feed Feed
Circuit
Circuit of Turns Per Ash Rate Solids Volume
Number
Notation Spirals Spiral (%) (TPH) (%) (GPM)
1 1 +2 2 4-3 39.9 2.52 24.1 37.9
2T0 CT1
2 1 +2 +3 +4 4 4-3-4-3 42.1 2.66 24.2 39.2
2T0 3T0 4T0 CT1
3 1 +2 +3 +4 4 4-3-4-3 37.8 2.47 24.0 37.7
2T0 3T1 4T0 CT1
4 1 +2 +3 3 3-4-3 39.6 2.53 24.0 38.2
2T0 3T0 CT1
5 1 +2 +3 3 2-2-4 40.2 2.68 24.4 39.8
2T0 3T0 CT1
4.3.4.1 Spiral Circuit 1 (1 +2 Circuitry)
2T0 CT1
First experimental run was carried out on the existing spiral circuitry used at the Cardinal
coal preparation plant. The 1 +2 spiral circuitry employed a 4 turn primary spiral followed
2T0 CT1
by a 3 turn secondary spiral, both connected to the same central column. An auxiliary repulping
box along with a refuse cutter was installed at the fourth turn of first spiral. The refuse cutter was
set to a distance of 9 inches measured from the outside of the central column. Feed slurry was
introduced at the top of the first spiral unit and, after four turns, primary refuse was separated
using a refuse cutter that passed the refuse down through the central column. The remaining
slurry was remixed using a repulping box installed following the three-turn spiral unit. After
passing the secondary spiral, six timed product samples were collected simultaneously with the
111 |
Virginia Tech | help of a specially-designed product collection box (Figures 4.2 and 4.6). The combined products
SC1 and SC2 were clean products, while PT1 and PT2 were the tailings and SM3, SM4 and SM5
were the middlings. Clean coal form the second spiral was the final clean product, while refuse
from both spirals were combined and rejected. Middlings from the second spiral was recycled
back to the spiral feed. Figure 4.7(a) is a schematic illustration of this spiral circuit. The purpose
of testing this circuit was to determine the separation performance of the existing compound
spiral and to compare its separation efficiency with that of other four spiral circuits.
4.3.4.2 Spiral Circuit 2 (1 +2 +3 +4 Circuitry)
2T0 3T0 4T0 CT1
The second spiral circuit consisted of four short spirals that employed a 1 + 2 + 3
2T0 3T0 4T0
+ 4 circuitry. The first and third spirals were four-turn spirals, whereas the other spirals in the
CT1
circuit had three turns each. Figure 4.7(b) shows the flowsheet of this circuit. As shown, feed
was introduced at the top of the first spiral, which flowed by gravity through the rest of the spiral
circuit. At each spiral unit, refuse was separated and taken out of the circuit, while the remaining
slurry was repulped before being fed to the next spiral. The clean coal stream collected from the
fourth and final spiral was taken as the final clean product and the combined refuse from all the
four spirals taken as the discard stream. The middlings from the fourth spiral were recycled back
to the first spiral feed.
4.3.4.3 Spiral Circuit 3 (1 +2 +3 +4 Circuitry)
2T0 3T1 4T0 CT1
The third set of spiral experiments were conducted using a 1 + 2 + 3 + 4 spiral
2T0 3T1 4T0 CT1
circuit (see Figure 4.7(c)). This circuitry was similar to the second spiral circuit in terms of the
number and turns per spiral. The only difference between the two circuits was the separation of
the second spiral middlings stream in this circuit. After passing the first spiral, refuse was
112 |
Virginia Tech | removed and the clean was repulped and retreated on the second spiral. The second spiral
produced three distinct products, i.e., clean, refuse and middlings, by using appropriate product
splitters. The second refuse was removed and clean was remixed with an auxiliary repulper and
fed to the third spiral. Middlings from second spiral was permitted to flow by gravity back to the
feed sump of the spiral circuit. Refuse from the third spiral unit was also separated and taken out
of the circuit, while the clean stream was remixed before being fed to the fourth stage spiral. The
final clean product was collected after the fourth spiral. The refuse products from all the four
spirals were combined and discarded. Middlings from the fourth spiral were combined with the
second spiral middlings and both were recycled back to the feed of spiral circuit. The purpose of
testing this circuit was to assess the effect of recycling second stage spiral middlings on the
overall separation efficiency of the second spiral circuit.
4.3.4.4 Spiral Circuit 4 (1 +2 +3 Circuitry)
2T0 3T0 CT1
The fourth set of experiment runs was performed using a circuit that consisted of three
short spirals. This 1 + 2 + 3 circuitry used 3, 4 and 3 turns per spiral, respectively. The
2T0 3T0 CT1
circuit layout is shown in Figure 4.8(a). The first spiral was fed fresh spiral circuit feed and the
first and second spiral clean was remixed and repulped before being fed to the next spiral unit.
The final clean product was produced by the third spiral unit, while the refuse from all of the
three spirals were combined and discarded. The third spiral middling was sent back to the spiral
circuit feed sump.
113 |
Virginia Tech | 4.3.4.5 Spiral Circuit 5 (1 +2 +3 Circuitry)
2T0 3T0 CT1
Finally, the fifth experiment was performed using a 1 + 2 + 3 spiral circuit.
2T0 3T0 CT1
Figure 4.8(b) shows the flowsheet for this spiral circuit. Both circuit configurations used in
experiment numbers 4 and 5 were identical in terms of their layout, but differed in terms of the
number of turns per spiral. Both circuits consisted of three spirals, but the fifth circuit had only
two turns per spiral for the first two spiral units. Three turns were used on the third spiral of this
circuit. Slurry feed was introduced at the top of the first spiral and, after passing the first spiral,
refuse was separated and the remaining slurry was remixed and retreated on the second spiral
unit. Again, the second spiral refuse was separated and the remaining slurry was repulped and
rewashed on the third spiral unit. Clean product from the third spiral was taken as final clean
product and the middlings were recycled back to the feed of spiral circuit. Refuse streams from
all the three individual spiral units were discarded. The purpose of testing spiral circuits number
4 and 5 were to determine the optimum number of turns before repulping the slurry.
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Virginia Tech | 4.3.5 Procedure
For all testing work, the coal feed consisted of nominal 1 x 0.15 mm size particles. As
shown in Figure 4.5, the minus 1 mm fraction of raw coal that passes through the raw coal
deslime screen was pumped to a bank of 15-inch diameter raw coal classifying cyclones. The
cyclone underflow (1 x 0.15 mm) was directed to the spiral feed sump where it was diluted to the
correct percentage of solids before being fed to the spiral circuitry. The spirals feed slurry was
first introduced to a distributor that over flowed into six sets of triple-start compound spirals.
One feed line from the distributor was used to feed the experimental spiral circuits used in this
study. A sampling port was also provided in the feed line so that a representative feed sample
could be collected. After passing through the spiral circuit the slurry stream was diverted into
appropriate sample containers and timed samples of all the products were collected. A total of 46
samples were collected during this test program (i.e., 8 samples form first test, 10 from second
test, 11 from third test, 9 from fourth test and 8 samples from fifth test). Sampling points are
shown by diamond symbols in Figure 4.7 and Figure 4.8. After weighting the feed and product
slurry samples, the products were subjected to wet-sieve analysis. Each sample was partitioned
into seven distinct size classes, i.e., +1 mm, 1 x 0.6 mm, 0.6 x 0.3 mm, 0.3 x 0.15 mm, 0.15 x
0.074 mm, 0.074 x 0.044 mm and -0.044 mm. Solids from each size class were filtered, dried,
weighted and analyzed for ash contents. During the whole experimental testing program, a total
of 322 (46 x 7) coal samples were collected, prepared and analyzed. ASTM standards were
followed throughout the experimental, sampling and ash analysis work.
Table 4.1 shows the design and operating parameters used in the spiral testing
experiments. The test data obtained from the research work, which included feed rate, percent
solids, volume flow rate, particle sizing and ash analyses, was adjusted using a spreadsheet based
117 |
Virginia Tech | mass balance routine. Variations in the coal feed characteristics made it difficult to calculate and
compare the performance of different spiral circuits. To overcome this problem, all the
calculations such as separation efficiency, overall yield, combustible recovery and organic
efficiencies were performed on balanced and normalized data.
4.4 Results and Discussion
4.4.1 Separation Efficiency
In the current research work, separation efficiency curves are used to compare the
cleaning performances of the various spiral circuits examined in this study. The concept of
separation efficiency has been discussed by many authors and is defined as the theoretical
percentage of feed material that passes through an ideal separation (Stevenes and Collins, 1961;
Schultz, 1970; Salama, 2001). As described earlier, a separator that results in a zero percent
selective separation is represented in recovery rejection plots by a dashed diagonal line passing
between the top-left corner and bottom-right corner (Figure 4.9 to Figure 4.13). In other words,
the separation efficiency line joins the 100% recovery and 0% ash rejection point with the 0%
combustible recovery and 100% ash rejection point. Likewise, a perfect separation is represented
by a single point on the top-right corner of the plot where both the combustible recovery and ash
rejection is 100%. In fact, these two boundaries represent the separation efficiencies of 0% and
100%, respectively. Other separation efficiency values are represented by the lines parallel to the
diagonal as shown dashed parallel lines in the recovery rejection plots. Each point along the
curve can be represented by a single value of separation efficiency, which reflects the trade-off
between recovering combustibles and rejecting ash. The optimum separation efficiency is
obtained for those combinations of operating conditions that give data points in the right-upper
most elbow of the performance curve.
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Virginia Tech | 4.4.2 Comparison of Separation Efficiencies
Figures 4.9 to 4.13 represent the size-by-size recovery rejection curves for all the five tested
spiral circuits. In general, similar trends have been found in separation efficiencies for all of
these spiral circuit configurations. The main points to be noted include:
• In each circuit, the highest separation efficiency was achieved by a coal feed in the size
range of 0.6 x 0.150 mm.
• The separation performance starts decreasing for feed particle sizes either greater than 0.6
mm or less than 0.150 mm.
• A significant deterioration in the performance is noted for the coal particles finer than
0.074 mm.
• The maximum separation efficiency for all feed size classes was achieved by spiral
circuit two (i.e., 1 + 2 + 3 + 4 ) and spiral circuit three (1 + 2 + 3 + 4 ),
2T0 3T0 4T0 CT1 2T0 3T1 4T0 CT1
followed by spiral circuit one (1 + 2 ), spiral circuit four (1 + 2 + 3 ) and
2T0 CT0 2T0 3T0 CT1
spiral circuit five (1 + 2 + 3 ), respectively.
2T0 3T0 CT1
• For a coal feed in the size range of 0.6 x 0.15 mm, both spiral circuit two (1 + 2 +
2T0 3T0
3 + 4 ) and spiral circuit three (1 + 2 + 3 + 4 ) were capable of achieving a
4T0 CT1 2T0 3T1 4T0 CT1
separation efficiency of at least 80%.
• The maximum separation efficiency for both spiral circuit one (1 + 2 ) and spiral
2T0 CT0
circuit four (1 + 2 + 3 ) was nearly identical at approximately 77%.
2T0 3T0 CT1
• For the same feed of size class of 0.6 x 0.150 mm, spiral circuit five (1 + 2 + 3 )
2T0 3T0 CT1
was the least efficient in terms of separation efficiency (approximately 70%).
• The separation efficiency for coal feeds in the size range of +1 mm and 1 x 0.6 mm was
the highest (approximately 70 to 72%) in the second, third and fourth spiral circuits,
119 |
Virginia Tech | followed by spiral circuit one (approximately 68 to 72%) and spiral circuit five
(approximately 65%).
• Spiral circuit two (1 + 2 + 3 + 4 ) and spiral circuit three (1 + 2 + 3 +
2T0 3T0 4T0 CT1 2T0 3T1 4T0
4 ) were the only circuits that were capable of achieving a high separation efficiency of
CT1
63 to 65% for coal feeds in the particle size range of 0.150 x 0.074 mm.
A comparative study for the results plotted in Figures 4.9 to 4.13 indicate that, under the
same operating conditions, separation efficiency of a spiral circuit depends on:
• the number of cleaning stages in a spiral circuit,
• the number of turns per spiral before repulping, and
• the recycling of middling streams.
These results suggest that highest separation efficiency for all feed size classes can be achieved
either by spiral circuit two (1 + 2 + 3 + 4 ) or spiral circuit three (1 + 2 1 + 3 +
2T0 3T0 4T0 CT1 2T0 3T 4T0
4 ). It was also concluded from the comparison between separation efficiencies of spiral circuit
CT1
four (1 + 2 + 3 ) and spiral circuit five (1 + 2 + 3 ) that repulping after just two
2T0 3T0 CT1 2T0 3T0 CT1
turns was insufficient to achieve a good separation efficiency and at least three turns were
required before repulping to maintain a high separation performance.
120 |
Virginia Tech | It is interesting to note that in all spiral circuit configurations shown in Figure 4.14, the
overall separation efficiency for the products collected at splitter position “SC1” was lower than
that obtained for position “SC2”. In other words, the proportion of misplaced rock particles in
the cleaner product “SC1” is more than the same in the product “SC2”.
One possible explanation for this trend may be that the ultrafine high-ash particles (minus
0.074 mm) in the spiral feed usually tends to report with the water. The data obtained by the size-
by-size product analysis at different splitter positions indicates that, on average, the proportion of
high-ash ultrafine particles (minus 0.074 mm particles containing 64.4% ash) in the “SC1” clean
product was 16.97% compared with that of 9.91% in the product collected at the “SC2” splitter
position. Thus, the apparently lower separation efficiency at splitter position “SC1” seems to be
mainly attributed to the presence of a higher percentage of ultrafine high-ash particles. However,
when the size-by-size separation efficiencies were plotted against different splitter positions in
Figure 4.15, the data show a similar trend for all of the coarser size fractions (+1 mm, 1 x 0.6
mm, 0.6 x 0.3 mm and 0.3 x 0.150 mm) as well.
The trend of high separation efficiency at splitter position “SC2” can better be explained
using the separation mechanism presented by Luttrell et al. (2007). This work identified two
counter-rotating flows across the spiral profile that converges along an imaginary line of
separation (Figure 4.16). The counter-clockwise flow in the lower rotation zone moves the
lighter particles towards the outer wall of the spiral and the heavier particles settle down and are
carried to the inner side of the spiral for rejection. Clockwise rotation is responsible for providing
a good refuse product that was relatively free of coal, while the counter-clockwise flow in the
upper rotating section stratifies the pure coal particles along the outer wall. Dense rock particles
that are entrapped in the upper zone tend to settle against the wall and are pinned there by the
124 |
Virginia Tech | upward counter-clockwise flow (Luttrell et al., 2000, 2007). Eventually, these entrapped high-
density particles report with the clean coal and, as such, are probably responsible for lower
separation efficiency at splitter position “SC1”.
Particle separation behavior across the spiral trough has been a continuous source of
confusion in the literature. A number of researchers have tried to explain the particle separation
mechanism including Holland-Batt (1990, 1992, 1994, and 1998), Richards and Palmer (1997),
Kapur and Meloy (1998) and Luttrell et al. (1998). Figure 4.16 helps to understand the flow
pattern and particle behavior across the spiral trough. In this figure, separation efficiency data
has been plotted against the distance from outer wall of the spiral trough. The plot was then
superimposed on the schematic diagram showing the particle separation and slurry flow patterns
across the spiral trough. The highest separation efficiency at the splitter position “SC2” and
lower separation efficiencies on either side of the “SC2” provide striking evidence for the
presence of two counter-rotating flows. As shown in Figure 4.16, the dashed line (at a distance of
8.65 cm measured from inside of the outer wall of the spiral trough) that passes through the
peaks of the separation efficiency curves is the line of separation which marks the boundary
between the counter-rotating flows on the spiral trough.
Frequent repulping of the feed slurry deteriorates the overall and size-by-size separation
efficiencies of spiral circuit 5 as shown in Figure 4.14 and Figure 4.15(E) respectively. A close
examination of the size-by-size separation efficiency curves shown in Figure 4.15(E) indicates
that the repulping after two turns helps the entrapped coarser (+1 and 1 x 0.6 mm) particles to
escape from the upper zone of high velocity flow. However, frequent repulping of slurry does
not provide the time necessary for the particles to segregate themselves into proper reject or
clean coal streams and, hence, results in a lower overall separation efficiency of the spiral circuit.
125 |
Virginia Tech | yield dropped significantly to 59% as shown by Figure 4.17(B). The best overall performance
was achieved by spiral circuit three (1 + 2 + 3 + 4 ) with a clean coal yield of 66% and
2T0 3T1 4T0 CT1
an organic efficiency of 81.2%. As stated previously, the only difference between spiral circuit
two and spiral circuit three was that the second middlings stream in spiral circuit three was also
recleaned along with the final stage middlings of the last spiral. Surprisingly, the clean coal yield
and organic efficiency achieved by the spiral circuit four (1 + 2 + 3 ) was nearly the same
2T0 3T0 CT1
as that of achieved by spiral circuit one (1 + 2 ). Finally, the data in Figure 4.17(E) shows
2T0 CT1
that among all the tested spiral circuits, spiral circuit five (1 +2 +3 ) performed the poorest
2T0 3T0 CT1
in terms of clean coal yield and organic efficiency. The overall yield and organic efficiency
obtained from this circuit was 56.2 % and 70.8%, respectively. Figure 4.8 provided previously
showed that spiral circuit five (1 + 2 + 3 ) incorporated auxiliary repupers after every two
2T0 3T0 CT1
turns for the first two cleaning stages. Thus, it is concluded that repulping after two turns actually
destroyed the separation process and caused in a lower separation efficiency, a lower clean coal
yield and lower organic efficiency and as shown by Figure 4.17(E).
129 |
Virginia Tech | Spiral circuit three (1 + 2 + 3 + 4 ) not only provided the overall best performance, but
2T0 3T1 4T0 CT1
also proved to be superior in all the individual feed size classes as well. Figure 4.18 shows the
clean coal yield and ash curves for individual size classes for the five spiral circuits.
Size-by-size clean coal yield and organic efficiencies at a clean ash of 10% are tabulated
in Table 4.2 and Table 4.3 respectively. In both of these data tables, the highest clean coal yield
and organic efficiencies are highlighted in bold letters. In each spiral circuit, the lowest values
for clean coal yield and organic efficiency was obtained for a feed size of plus 1 mm and the
highest clean coal yield was obtained when treating the 0.3 x 0.150 mm size fraction. The
interesting point here is that in each spiral circuit, cumulative clean coal yield and organic
efficiencies improved with the decrease in feed size up to 0.150 mm.
Finally, as shown in Table 4.2, spiral circuit three (1 + 2 + 3 + 4 ) provided an
2T0 3T1 4T0 CT1
overall yield of 66%, while at the same clean coal ash of 10%, an overall yield value of 64% was
provided by the existing compound spiral circuit, i.e., spiral circuit one (1 + 2 ). In other
2T0 CT1
words, spiral circuit three (1 + 2 + 3 + 4 ) increased the yield by 1.9%, while
2T0 3T1 4T0 CT1
maintaining the same clean coal ash of 10% achieved by the existing spiral circuit. For example,
in a 100 tph spiral circuit, this net increase in the yield can be translated into a dollar value of
$751,830 per year (i.e., 1.9 ton/hr x $65.95/ton x 6,000 hrs/yr).
131 |
Virginia Tech | Table 4.2 Summary of size by size cumulative yield at 10% of product ash.
Cumulative yield (%) for different size classes (mm)
Spiral Circuits
+ 1 1 x 0.6 0.6 x 0.3 0.3 x 0.15 Overall
1 +2 44.2 55.1 67 70.4 64.1
2T0 CT1
1 +2 +3 +4 46.2 54.8 70.3 70 59.0
2T0 3T0 4T0 CT1
1 +2 +3 +4 49.8 60 72.2 73.9 66.0
2T0 3T1 4T0 CT1
1 +2 +3 48.4 54.8 70.3 72 64
2T0 3T0 CT1
1 +2 +3 43.1 51.8 64.9 68.4 56.2
2T0 3T0 CT1
Table 4.3 Summary of size-by-size organic efficiencies at 10% product ash.
Organic efficiency (%) for different size classes (mm)
Spiral Circuits
+ 1 1 x 0.6 0.6 x 0.3 0.3 x 0.15 Over All
1 +2 74.63 81.71 88.84 91.29 78.72
2T0 CT1
1 +2 +3 +4 77.61 82.56 92.03 90.94 72.66
2T0 3T0 4T0 CT1
1 +2 +3 +4 82.24 87.32 93.17 93.76 81.16
2T0 3T1 4T0 CT1
1 +2 +3 80.45 81.22 91.69 91.76 78.62
2T0 3T0 CT1
1 +2 +3 72.24 77.68 87.70 88.94 70.81
2T0 3T0 CT1
4.4.5 Combustible Recovery
Figure 4.19 shows the size-by-size combustible recovery and ash curves for the five
experimentally tested spiral circuits. The data shows that spiral circuit three (1 + 2 + 3 +
2T0 3T1 4T0
4 ) offered the best combustible recovery in all size classes, while the lowest combustible
CT1
recovery was obtained by using the spiral circuit five (1 + 2 + 3 ). Table 4.4 compares the
2T0 3T0 CT1
size-by-size and overall combustible recovery at 10% clean coal ash for the experiments
conducted on each of the five spiral circuits. Spiral circuit one (1 + 2 ) and spiral circuit four
2T0 CT1
(1 + 2 + 3 ) were found to have same overall combustible recovery, but differed in
2T0 3T0 CT1
combustible recoveries for individual feed size classes. As indicated by Figure 4.19, spiral
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Virginia Tech | Table 4.4 Comparison of spiral circuits (Summary of size by size combustible recovery at
the 10 % of clean coal ash).
Combustible recovery (%) for different size classes (mm)
Spiral Circuits
+ 1 1 x 0.6 0.6 x 0.3 0.3 x 0.15 Overall
1 +2 75.6 81.9 90.1 92.1 77.1
2T0 CT1
1 +2 +3 +4 79.2 82.5 95 91.5 71.5
2T0 3T0 4T0 CT1
1 +2 +3 +4 84 89.1 96.8 95.1 79.6
2T0 3T1 4T0 CT1
1 +2 +3 82 80.9 95 94 77.1
2T0 3T0 CT1
1 +2 +3 74.1 77 88 89 68.5
2T0 3T0 CT1
4.4.6 Spline Curves
To show the incremental changes in clean coal yield and combustible recovery, spline
curves were fit to the experimental data so as to improve numerical comparisons. The size-by-
size coal yield and combustible recovery spline curves are shown in Figure 4.20 and Figure 4.21,
respectively. These figures show the size-by-size data for the products collected at splitter
positions “SC1”, “SC2” and “SM3”. In the plots, dashed circles show the actual points for clean
coal yield and recovery for the products collected splitter position “SC1”. The important item to
note here is that spiral circuit three (1 + 2 1 + 3 + 4 ) is capable of producing both
2T0 3T 4T0 CT1
coking and thermal coal products. For example, the low ash (less than 5%) clean coal product
collected at splitter position “SC1” in spiral circuit three can be sent to a coking plant, while the
combined products collected at splitter positions “SC2” and “SM3” can still meet the ash
requirements for coal products used for power generations (Figure 4.20 and Figure 4.21).
Moreover, it can be clearly seen in the spline curves that amongst all the experimentally tested
spiral circuits, spiral circuit three provides the maximum clean coal yield and recovery at any
selected clean coal ash.
136 |
Virginia Tech | 4.5 Conclusions
A full-scale experimental study on five different spiral circuits was conducted at an
industrial coal preparation plant. The results suggest that a four-stage spiral described by the
shorthand notation “1 + 2 + 3 + 4 ” offered the best option for improved separation
2T0 3T1 4T0 CT1
efficiency, clean coal yield and combustible recovery. Preliminary calculations indicate that this
spiral circuit is capable of increasing clean coal yield by 1.9%, while maintaining the same ash
contents as achieved by the existing two-stage compound spiral circuitry (1 + 2 ) currently
2T0 CT1
installed in the plant. Moreover, spline curves fit through the experimental data indicate that the
new four-stage spiral circuit, when used with appropriate product splitter settings, can
simultaneously produce both low-ash coking coal and high-ash thermal coal products. Finally,
the data also suggest that repulping after two turns in a coal spiral circuit does not provide
sufficient residence time for good separations of coal and rock. On the other hand, repulping
after three or more turns is recommended to improve spiral separation performance.
139 |
Virginia Tech | CHAPTER 5 ENHANCED SULFUR REJECTION USING COMBINED
SPIRALS FLOTATION CIRCUITS
5.1 Abstract
A detailed study was conducted to evaluate the partitioning of pyrite within fine coal
circuits. The investigation, which included both laboratory and pilot-scale test programs,
indicated that density-based separations are generally effective in reducing sulfur due to the large
density difference between pyrite and coal. On the other hand, the data also showed that sulfur
rejections obtained in froth flotation are often poor due to the natural floatability of pyrite.
Unfortunately, engineering analyses showed that pyrite removal from the flotation feed using
density separators would be impractical due to the large volumetric flow of slurry that would
need to be treated. On the other hand, further analyses indicated that the preferential partitioning
of pyrite to the underflow streams of classifying cyclones and fine wire sieves could be exploited
to concentrate pyrite into low-volume secondary streams that could be treated in a cost effective
manner to remove pyrite prior to flotation. On the basis of this study, an enhanced sulfur
rejection circuit was designed and implemented on an industrial scale in the Illinois coal basin.
This paper describes the rationale for the design of this new fine coal cleaning circuit and
presents the data obtained from the full-scale sampling program.
5.2 Introduction
There has been an increasing worldwide interest to reduce overall atmospheric levels of
sulfur dioxide, the primary precursor to acid rain. Environmental legislation to reduce sulfur
emissions to the atmosphere have been already in place in many countries and are under
consideration in most of other countries. In United States, the Clean Coal Act amended in 1990
142 |
Virginia Tech | puts strict restrictions and set targets for overall sulfur dioxide emissions to the atmosphere. Coal
when burned produces pollutants such as sulfur oxides, nitrogen oxides and particulate matter.
The total United States SO2 emissions from coal utilization are approximately 15.9 million tons
per year. This level includes emissions arising from power generation of 10.9 million tons per
year. By 2014, the EPA aims to reduce SO2 emissions from coal fired power plants by 71
percent from 2005 levels (US-EPA, 2012). To meet the air pollution standards, many coal fired
power plants in United States are switching from high sulfur eastern and midwestern coal
feedstocks to low sulfur western coals. However, the coal from western regions has increased the
transportation cost for those power plants operated in eastern and midwestern regions.
Sulfur contents and type in coals are highly variable and depends upon the depositional
environment of the coal. Total sulfur contents in coal may vary from a low of 0.5% sulfur to as
high as 11% (Monticello and Finnerty, 1985).There are a number of technologies available
throughout the coal life cycle that can be used to reduce the sulfur in coal. These include
(Cavallaro et al., 1991; Ohtsuka, 2009):
• Physical removal of sulfur from coal before combustion
• Magnetic and electrostatic separation of sulfur form coal
• Chemical cleaning of high sulfur coals
• Biological cleaning of high sulfur coals
• Conversion of high sulfur coals to a low sulfur clean fuel by gasification or liquefaction
process
• Blending of high sulfur coals with low sulfur coal feeds
• In bed desulfurization
• Flue gas desulfurization
143 |
Virginia Tech | • Wet scrubbers
In the past, coal cleaning operations focused only on reducing the percentage of ash
forming minerals from coal and not necessarily targeted to reduce the sulfur contents. However,
the past decade witnessed a continuous increase in the research and development of physical
sulfur removal processes using coal beneficiation methods (Celik and Yildirin, 2000; Kawatra
and Eisele, 2001; Rubiera et al., 1997; Mohanty et al., 2008; Shah et al., 2001; Mbamba et al.,
2012). As a result of these studies, it appears that a significant amount of sulfur can be removed
from high sulfur coals during coal processing operations. Moreover, removal of sulfur from high
sulfur coal during processing operations offered many added advantages such as reducing the
load on flue gas scrubbers, reduced transportation cost and simultaneous removal of ash and
sulfur bearing minerals from coal.
Physical coal processing methods generally exploit either the difference between specific
gravities or surface properties of coal and rock forming minerals. In the United States, modern
coal preparation plants may include as many as four separate processing circuits for cleaning the
coarse (+10 mm), small (10 x 1 mm), fine (1 x 0.15 mm) and ultrafine (-0.15 mm) coal feeds..
Coarse and small coal fractions are almost exclusively cleaned using dense medium separation
processes. A wide variety of viable cleaning circuit configurations exist to treat fine (-1 mm)
coal, such as water-based density separators (spirals or water-only cyclones) as well as various
types of surface-based separators (conventional or column flotation processes) (Luttrell et al.,
2007).
An important problem associated with the cleaning of high sulfur coal is the inferior
sulfur rejection performance of fine coal cleaning circuitry (Mohanty et al., 2008). On one hand,
density separation devices can reduce clean coal sulfur levels within the limits set by the
144 |
Virginia Tech | washibility characteristics of high sulfur fine coal and the limit imposed by the specific gravity
cut point associated with the fine coal cleaning method. On the other hand, froth flotation faces
another challenge of recovering a significant amount of pyrite when treating high sulfur ultrafine
coal. Although coal pyrite particles are well liberated at particle sizes down to nominal flotation
feed sizes (minus 0.150 mm), unwanted mineral matter often tends to report to the clean coal
product of a flotation cell due to entrainment and the unpredictable hydrophobic nature of coal
pyrite (Kawatra and Eisele, 2001). Thus, froth flotation, which otherwise provides excellent ash
separation performance, often performs poorly in in terms of sulfur rejection. As a matter of fact,
in a number of laboratory flotation experiments conducted during current research work, sulfur
contents of the froth product were often found to be higher in sulfur than the flotation feed.
In order to address the challenge posed by high sulfur fine coal, several research studies
have been conducted in an attempt to solve this problem (Honaker et al., 1995; Luttrell et al.,
1995; Mcalister, 1998; Mohanty and Honaker, 1999; Celik and Yildirin, 2000; Shah et al., 2001;
Kawatra and Eisele, 2001; Rubiera et al., 1997; Mohanty et al., 2008; Mbamba et al., 2012).
Most of these studies examined and compared the coal pyrite separation performance of
individual fine coal cleaning technology. For example, Mohanty and Honaker (1995, 1999)
found that for a feed size of 0.6 x 0.045 mm, sulfur rejection performance of enhanced gravity
separators was better as compared to that of a flotation process. Kawatra and Eisele (2001)
conducted a detailed investigation of coal desulfurization using different coal preparation
methods. Unfortunately, pyrite recovery of the fine coal circuits typically employed in coal
processing facility is not as efficient as their ash removing efficiency. In fact, there is no
standardized method that exists in the coal preparation industry for sulfur removal. Thus, there is
145 |
Virginia Tech | a need for an efficient, economical and industrially acceptable method for pyrite removal from
high sulfur fine coal.
Hydrophobic coal pyrite particles (partially oxidized pyrite particles are often
hydrophobic) or partially liberated pyrite often behave as floatable particles and report to the
froth product. The same pyrite particles can be rejected by a density based separator because of
their relatively high density. Unfortunately, density separators are not effective in rejection
ultrafine particles such as clays, which contribute significantly to raising the ash content and
lowering the heating value of the final product. In conclusion, it has been found that different
types of fine coal separators are effective in removing a particular type of mineral. Thus, a
combination of two different fine coal separation processes often gives better ash and pyrite
separation efficiency as compared to that of multiple stages of a single type of separation process
(Kawatra and Eisele, 2001).
In view of this fact, an extensive laboratory and pilot-scale experimental study was
conducted at Virginia Tech to identify optimum methods for simultaneously increasing ash and
sulfur rejections from fine coal. The objective of this investigation was to evaluate the coal pyrite
separation performance of each fine coal separation process used in this study and to design a
fine coal circuitry that gives highest performance for both ash and pyrite rejection. Unit
operations examined in this study included a water-only cyclone, two stage compound spiral and
froth flotation cell. On the basis of experimental findings, an enhanced sulfur rejection fine coal
circuitry was developed. The new circuitry combined both density-based (spirals) and surface-
based (flotation) separation processes. It was found that the combined spiral and flotation circuit,
when properly located behind various types of sizing and classification units, achieved the
highest ash and sulfur rejection performance amongst all individual unit tested.
146 |
Virginia Tech | 5.3 Experimental
5.3.1 Plant Description
Figure 5.1 shows a generic flowsheet of a coal processing plant being investigated in this
research. It is located in midwestern region of United States and used to treat 3" x 0 mm coal
feed. The plant feed is fed to a double deck banana screen which is used as a combined raw coal
and deslime screen, cutting at 1/2 inch on the top deck and 1 mm on the bottom deck. The coarse
(3" x 1 mm) size fraction reports to a bank of dense medium cyclones for density separation.
Undersize (minus 1 mm) of raw coal banana deslime screen is classified at a cut size of 0.150
mm by using a bank of 15 inch diameter raw coal classifying cyclones. Underflow from raw coal
classifying cyclones (1 x 0.150 mm) flows to a bank of triple start compound spirals. Clean from
spirals reports to clean coal sieve bend screens. Undersize (0.35 x 0) of clean coal sieve bend
along with raw coal classifying overflow reports to a bank of deslime cyclones. Finally, froth
flotation is used to clean deslimed fine coal (0.150 x 0.044 mm) stream. Combined deslimed
spiral and flotation concentrate is dewatered by screen bowl centrifuges. A single thickener is
used to treat plant fine tailings. Thickener underflow is pumped to an impoundment and the
coarse refuse is hauled to a coarse refuse area. Prior to this study, the plant management had
expressed concerns regarding the poor sulfur rejections achieved in the fine coal circuitry.
147 |
Virginia Tech | Table 5.1 Size by size average ash and sulfur analysis for
the undersize stream of clean coal sieve bend.
Feed Feed Feed Feed
Size Fraction Ash Sulfur
(mm) (%) (%) (%)
Plus 0.25 4.59 11.54 3.14
0.25 x 0.15 21.52 9.52 3.05
0.15 x 0.075 29.29 21.49 4.01
0.075 x 0.044 11.13 55.34 8.56
Minus 0.044 33.46 74.07 7.83
Total 100 39.82 5.55
Thus, the high sulfur ultrafine clean coal sieve undersize stream was selected as the feed
slurry for this pilot-scale investigation. In order to normalize the fluctuations in the feed
characteristics, a total of six barrels of coal slurry samples were collected over a period of 24
hours. Table 5.1 shows that the average total sulfur content of the collected feed sample was
5.5%, thus this coal sample can be classified as a high sulfur coal. Table 5.1 also illustrates
another interesting aspect of the fine coal feed, i.e., the majority of the ash and sulfur is
concentrated in the finer fractions of the feed.
5.3.3 Procedure
Following fine coal cleaning circuitries were experimentally tested for their sulfur and
ash separation performance:
• Spirals only circuit
• Water only cyclone circuit
• Froth flotation circuit
• Combined spirals and flotation circuit
149 |
Virginia Tech | Figure 5.2 provides a schematic of pilot-scale circuitry. As mentioned earlier, coal slurry
from undersize clean coal sieve bend screen was used to prepare feed coal slurry of desired
percent solids into a 240 gallon capacity feed sump. For all the experimental tests, coal feed
consisted of 0.25 x 0.044 mm sized particles. During spiral and water-only cyclone testing, the
slurry was held in suspension using a 25 cm diameter blade mixer. Slurry from the sump was
pumped at a controlled rate using a variable-speed centrifugal pump equipped with a 35-cm
diameter impeller. Feed slurry from the pump was passed up to an upper level floor to each unit
operation being testing. A sample port was provided in the vertical feed line so that a
representative feed sample could be collected. After passing through the cleaning unit, timed
samples of the products were collected by diverting the full flow of the product streams into
sample containers. For water-only cyclone tests, the collected products included timed samples
of overflow and underflow products. Likewise, the two-stage compound spiral included a
product box partitioned to collect six different samples across the profile of the second stage
spiral as well as an upper draw-off point for the collection of primary refuse from the first stage
spiral. This configuration made it possible to simultaneously collect timed samples of clean coal,
refuse and five different middlings products so that complete recovery-rejection curves could be
generated for each spiral test run. For the flotation evaluation, timed kinetic tests and release
analysis tests were conducted on the same feed coal. In the combined spiral-flotation circuit,
spiral clean coal product was used as feed to the froth flotation cell. After collecting and
weighing the slurry samples, the solids were filtered, dried, weighted and analyzed for size-by-
size sulfur and ash contents. The experimental data were then evaluated using a spreadsheet-
150 |
Virginia Tech | Product collected at splitter position “P1” represents the cleanest product taken at the
outer most position across the spiral profile, while “P6” represents the high density reject product
taken at the inner most position near the center support tube for the spiral. Primary spiral refuse
“P7” was separated by the primary refuse cutter and collected through the central column.
Throughout this research study, combined products of “P1” and “P2” comprise the clean coal
while products from “P3” to “P5” are the middlings. The combined products collected at splitter
position of “P6” and “P7” are the spiral circuit reject. Table 5.2 shows the operating parameters
for different spiral tests used in this study. These parameters were selected in accordance with
the recommendations and guidelines provided by Honaker for cleaning of ultrafine coal using
spirals (Honaker et al., 2007). Spiral test 1 was conducted using a relative high mass feed rate
(1.04 TPH/start), feed percent solids (19.72%), and volumetric feed rate (19.43 GPM) whereas
spiral test 5 was conducted on lowest values of these operating parameters. Operating parameters
for other spiral tests were in between these two extremes.
Table 5.2 Operating parameters for spiral only circuits.
Spiral Feed Feed Feed
Test Rate Solids Volume
(Number) (TPH) (%) (GPM)
1 1.04 19.72 19.43
2 1.46 19.73 27.35
3 0.80 12.44 24.48
4 0.65 12.20 20.21
5 0.46 12.15 14.56
Figure 5.4 and Figure 5.5 elucidate the cumulative effect of splitter positions across the
spiral profile on ash and sulfur separation efficiencies. Spirals are typically utilized to clean 1 x
153 |
Virginia Tech | 0.15 mm coal feed. Thus, coal spirals used to treat ultrafine (0.25 x 0 mm) coal yield poorer ash
and sulfur separation efficiencies as indicated by the plots. Other important points to be noted in
the figures are as follows.
• Maximum ash and sulfur separation efficiencies were approximately 17% and 27%,
respectively.
• Among all the tests, the maximum ash and sulfur separation efficiencies were achieved
during spiral test 1.
• At the same splitter position, sulfur separation efficiencies for each spiral test are better
when compared to the corresponding ash separation efficiencies for the same spiral test.
• The specific gravity of sulfur-bearing particles is higher as compared to that of ash-
bearing particles. Consequently, better sulfur separation performance may be due to the
high specific gravity of ultrafine sulfur, which helps these particles to segregate
themselves into reject streams.
• An interesting point to note is that both ash and sulfur separation efficiencies for all spiral
tests are minimum for the product collected at splitter position “P1”. In other words,
product “P1” contains a relatively high percentage of misplaced particles.
• For all the spiral tests, both the ash and sulfur separation efficiencies are comparatively
low for the products collected at splitter position “P1”.
• For all the spiral tests, both ash and sulfur separation efficiencies improve when products
from splitter positions “P2”, “P3” and “P4” are added to the clean product of “P1”.
154 |
Virginia Tech | Table 5.3 shows the size-by-size ash and sulfur reductions achieved by all spiral tests. In
order to have a fair performance analysis, all the spiral tests are compared at the same feed
assays. The data presented in Table 5.3 indicates that significant ash and total sulfur reductions
were achieved for all feed sizes except the minus 0.044 mm size class. For the plus 0.075 mm
feed coal, higher percentage reductions in product ash were noticed as compared to the minus
0.075 mm size class. In fact, feed size classes above 0.044 mm shows substantial decreases in
percent sulfur values for the products.
Overall separation performance achieved by different spiral circuit is presented in Table
5.4. The highest combustible recovery of 83.7% was achieved during the second spiral test,
while the highest sulfur and ash rejection was obtained during the first test run. For the overall
coal feed (0.25 x 0 mm), a maximum of 36.60% of the ash and 45.72% of the sulfur was rejected
(Table 5.4, Spiral test 1), while 50.77% ash and 49.85% sulfur rejection was obtained for the
plus 0.044 mm coal (Table 5.5, Spiral test 1). Because of its poor separation performance, the
minus 0.044 mm product size fractions were screened and removed.
The data provided in Table 5.5 represents the assays obtained from 0.25 x 0.044 mm feed
size fractions for all individual spiral tests. A comparison between Table 5.4 and Table 5.5
indicates that significant improvements in separation performance were achieved by removing
minus 0.044 mm size particles. For example, the combustible recovery of the 0.25 x 0.044 feed
size fraction for spiral test 1 improved by 7.97% when compared to the combustible recovery
achieved for a feed size of 0.25 x 0 mm for the same spiral test. Similarly, the clean coal ash and
sulfur values were reduced by 20.26% and 5.26%, respectively. Thus, to achieve an acceptable
sulfur rejection, the minus 0.044 size fraction needs to be removed from spiral feeds.
156 |
Virginia Tech | Figure 5.11 through Figure 5.15 represent the size-by-size combustible recovery and
sulfur rejection curves for all the spiral tests. The important point to note is that unlike the ash
separation efficiencies, good sulfur separation efficiencies were obtained for the ultrafine (minus
0.15 mm) coals. Moreover, the sulfur separation efficiencies for minus 0.15 mm coal feed are
better when compared to that of the plus 0.15 mm coal feed (Figure 5.11). The plots also show
that spirals perform better in terms of sulfur rejection as compared to ash rejection for minus
0.15 coal feeds. Other points to be noted from the Figure 5.11 to Figure 5.15 are as follows.
• The maximum separation efficiency of sulfur was obtained by maintaining medium feed
flow rates (i.e., 1 tph and 20 gpm).
• In contrast with the ash separation efficiency (Figures 5.6 to 5.10), plus 0.25 mm coal
feeds perform less well in terms of sulfur separation performance.
• The maximum sulfur separation efficiency was obtained for the 0.075 x 0.044 mm feed
size fraction, followed by the 0.15 x 0.075 mm feed size fractions.
• Even the minus 0.044 mm feed size fraction shows some degree of sulfur separation.
162 |
Virginia Tech | Table 5.6 Ash and total sulfur reductions achieved from the treatment of high sulfur 0.25 coals by
water-only cyclone (Feed ash = 39.82%, Feed sulfur = 5.55%).
WOC Feed Ash Sulfur Ash Sulfur Cumulative Combustible
Test Solids Product Tailings Product Tailings Rejection Rejection Yield Recovery
(Number) (%) (%) (%) (%) (%) (%) (%) (%) (%)
1 9 59.51 36.44 2.72 6.47 68.74 89.53 21.78 15.06
2 12 60.87 35.01 2.80 6.22 68.16 68.16 21.18 13.92
3 15 61.86 34.98 2.60 6.39 71.46 71.46 18.42 11.70
Figure 5.16 and Figure 5.17 summarize the size-by-size ash and sulfur separation
performance, respectively, for the water-only cyclone tests. In general, very poor ash separation
efficiencies were obtained for all feed size fractions. Sulfur separation efficiencies for the same
feed size are comparatively better than those of ash separation efficiencies at the same feed
percent solids (Figure 5.17). Interestingly, amongst all coal feed size fractions, the minus 0.044
mm coals attained maximum sulfur separation efficiency (25 to 30%). Poor separation of both
ash and sulfur results may be because the water-only cyclones are optimized to treat coal feed in
the size range of 3 x 1 mm.
Figure 5.18 graphically represents the size-by-size sulfur separation efficiencies versus
different feed percent solids of the water-only cyclone circuit. It shows that separation
performance of each size class increases with increase in feed percent solids and as such
minimum sulfur separation was obtained at 9% feed solids, while maximum separation was
achieved when the water-only cyclone was operated at 15% feed solids. In conclusion, the water-
only cyclone, when used to clean fine (minus 0.15 mm) high sulfur coal, tends to reject the
majority of the sulfur-bearing particles into the underflow reject stream, but the very low
combustible recovery of the unit made it impractical for cleaning ultrafine high sulfur coals.
166 |
Virginia Tech | 5.4.3 Froth Flotation Circuit
Froth flotation has proved to be an effective process for cleaning fine (minus 0.15 mm)
coals. In this research study, both release analysis and timed flotation tests were carried out using
a laboratory froth flotation machine (Denver Model D-12). The frother was Dowfroth 200 and
the fuel oil was used as a collector. In order to find out the natural floatability of the clean coal
product, flotation tests were also performed without the addition of any collector. In total, four
different flotation tests were carried out as shown by Table 5.7. The results show that, at a
comparatively lower yield of 49.75%, good ash and sulfur rejections were achieved by the
flotation circuit. The highest yield (66.32%) and combustible recovery (92.34%) corresponds to
the lowest ash (68.27%) and sulfur (32.83%) rejections.
Table 5.7 Ash and total sulfur reductions achieved by froth flotation circuit on high sulfur
minus 0.25 coal feed. (Feed ash = 42.62%, Feed sulfur = 5.92%).
Flotation Ash Sulfur Ash Sulfur Cumulative Combustible
Test Product Tailings Product Tailings Rejection Rejection Yield Recovery
(Type) (%) (%) (%) (%) (%) (%) (%) (%)
Release 14.85 70.40 5.64 6.31 82.72 53.07 49.75 74.01
Release* 14.80 84.26 5.28 6.48 78.88 44.61 60.39 89.19
Kinetic 19.55 82.94 5.84 6.27 70.62 37.83 63.83 89.27
Kinetic* 20.54 87.01 6.03 5.80 68.27 32.83 66.32 92.34
*Flotation tests with frother only
Size-by-size ash and sulfur separation performances achieved by the flotation-only tests
are shown in Figure 5.18 through Figure 5.22. The feed size fraction of 0.250 x 0.150 mm
showed the least ash separation efficiency, while the plus 0.25 mm coals showed a comparatively
168 |
Virginia Tech | 5.4.4 Combined Spirals and Flotation Circuit
Analysis of the sulfur separation performance achieved by each individual technology
tested in this study indicates that either the sulfur contents in the final clean coal concentrate
remains high (spiral or flotation-only circuits) or the combustible recovery of the clean coal
remains too low for an economical separation process (water-only cyclone circuit). However, the
data also suggests that the retreatment of spiral clean coal by a forth flotation process might
provide an improved sulfur rejection performance for ultrafine high-sulfur coal. Therefore, to
prove the concept, additional flotation experiments were conducted on the spiral clean product.
Initially, the high sulfur ultrafine coal feed sample was treated on a spiral. Afterwards, the spiral
clean product was retreated by the froth flotation process. Spiral refuse was rejected, while the
middlings were recycled back to the spiral feed. Ash and sulfur reductions achieved by the
combined spiral-flotation circuitry are illustrated as a schematic flow diagram in Figure 5.23.
An interesting point to be noted in Figure 5.23 is that the sulfur percentage of the final
product/froth concentrate (4.13%) is higher than that of flotation feed/spiral clean (4.1%). It may
be due to the fact that spirals tend to (i) remove only fully liberated high-density ultrafine pyrite
particles and (ii) concentrate both the partially liberated pyrite particles and the chemically bond
organic sulfur present in the coal feed. Moreover, as indicated by Figure 5.22, the froth flotation
process is not as efficient in removing sulfur-bearing minerals from coal as it is in removing ash-
bearing minerals. Therefore, froth flotation appears to concentrate the sulfur in the final clean
coal by removing more particles of ash-forming minerals than particles of coal pyrite.
172 |
Virginia Tech | which is clearly unacceptable. Although the flotation-only run provided good ash and sulfur
rejections, the clean coal yield and combustible recovery provided by the same circuit were
lower as compared to that of the spiral-only and combined spiral-flotation circuits. Amongst all
the coal cleaning alternatives tested during this work, the combined spiral-flotation circuitry
produces a coal with a lowest ash (10.77%) and sulfur (4.13%) at a reasonable combustible
recovery of 77.29%. In fact, a detailed comparison of the results (Table 5.8) indicates that the
combined spiral-flotation circuitry is superior in producing a low-ash and low-sulfur clean coal
product at a comparatively higher clean coal yield.
Table 5.8 Comparison of ash and total sulfur reductions achieved by different fine coal
cleaning circuits on high sulfur minus 0.25 coal feed.
Spiral WOC Flotation Spiral + FF*
Ash (%) 39.82 39.93 42.76 39.82
Feed
Sulfur (%) 5.55 5.69 5.98 5.55
Yield (%) 72.14 18.42 49.75 52.13
Recovery (%) 77.92 11.70 74.01 77.29
Clean
Coal
Ash (%) 35.0 61.86 14.85 10.77
Sulfur (%) 4.18 2.60 5.64 4.13
Yield (%) 13.45 81.58 50.25 33.46
Ash (%) 65.48 34.98 70.40 78.29
Reject Sulfur (%) 10.32 6.39 6.31 7.39
Ash Rejection (%) 36.60 71.46 82.72 65.79
Sulfur Rejection (%) 45.72 91.59 53.07 44.54
*Combined spiral and flotation circuit
The superiority of the combined spiral-flotation circuit is slightly masked by the fact that
the ash and sulfur contents of the flotation only circuit feed were slightly higher than those of for
174 |
Virginia Tech | the combined circuit feed. Thus, for a fair comparison, the performance of each circuit was also
compared based on separation efficiencies. The separation efficiencies were derived from
combustible recoveries versus ash/sulfur rejection plots.
Figure 5.24 and Figure 5.25 compares the size-by-size sulfur and ash separation
efficiencies of different fine coal cleaning alternatives evaluated in this research. The separation
efficiencies shown in these figures represent the best level of performance achieved by each coal
cleaning circuit used in this study. The combined spiral and flotation circuit was able to provide
the lowest clean coal sulfur contents over the entire range of particle sizes (Figure 5.24).
Although the sulfur separation efficiencies achieved by the spiral only circuit were nearly
comparable to the combined spiral-flotation circuit for the minus 0.15 mm feed coal. But the ash
separation efficiencies of spiral only circuit for the same feed sizes were less than that of
achieved by the combined spiral and flotation circuit (Figure 5.25). Ash separation efficiencies
of both flotation only and combined spiral and flotation circuits were the same for a feed size of
0.15 x 0.0.75 mm and for minus 0.075 mm coals flotation is the only process to achieve highest
ash separation performance (Figure 5.25). Amongst all the coal cleaning circuit experimentally
tested, the lowest ash and sulfur separation efficiencies were achieved by water-only cyclone
across all particle sizes. In conclusion, the combined spiral-flotation circuit is capable of
maintaining high sulfur and ash separation efficiencies across all particle size studied.
175 |
Virginia Tech | Spiral-only circuitry, when used to treat high sulfur ultrafine coal, provides good sulfur
separation efficiencies, but its clean product contains a high (35%) percentage of ash-forming
minerals. The water-only cyclone achieved high ash and sulfur rejections of 71.46% and 91.59%,
respectively. However, due to a poor clean coal yield of 18.42%, the separation efficiencies of
the water-only cyclone across all particles sizes tested remains very low. Finally, although the
froth flotation process provides the highest ash separation efficiencies for minus 0.15 mm coal,
the process fails to achieve the same levels of superior performance in terms of sulfur separation.
Based on the experimental data obtained from this research, a combined sieve screen,
spiral separator followed by a froth flotation process is recommended for the cleaning of high-
sulfur ultrafine (minus 0.25 mm) coal feeds. It is concluded that significant reductions both in
ash and sulfur contents of clean coal are possible by the combined sieve screen-spiral-flotation
circuitry, while maintaining a reasonable clean coal yield. Finally, a flowsheet is also proposed
for any coal preparation plant treating high sulfur ultrafine coal feeds.
5.8 Recommendations for Future Work
The main objective of this experimental work is to prove the concept that a combined
sieve screen, spiral and froth flotation circuit can practically achieve better sulfur separation
efficiency as compared to that of spirals, water-only cyclones or froth flotation only circuits.
Throughout the experimental work the same minus 0.25 mm high sulfur ultrafine coal feed used.
In order to estimate the exact improvements in ash and sulfur rejections offered by the combined
spiral and froth flotation circuit, it may be worth examining to experimentally test the same
combined circuit using ultrafine deslime (0.25 x 0.044 mm) high sulfur coal feed.
180 |
Virginia Tech | CHAPTER 6 ENGINEERING DEVELOPMENT OF THE
MICROSIEVE DRYING PROCESS
6.1 Abstract
The removal of moisture from fine coal has been a longstanding problem in the coal
preparation industry. While coal fines often represent as little as 10% of the total run-of-mine
feed, this size fraction may contain more than a third of the total moisture in the final marketed
product. Existing thermal dryers can effectively reduce moisture; however, these massive units
require very large capital expenditures and have become a target of increased environmental
scrutiny. Likewise, existing mechanical equipment for fine coal dewatering tend to produce
unacceptably high moistures that often cannot be tolerated on existing coal contracts. In light of
these issues, a mechanical, non-thermal patent-pending dewatering process has been developed
by NDT™. This article (i) reviews the working features of this novel process, (ii) presents
experimental results obtained from recent laboratory and pilot-scale test programs, and (iii)
discusses the potential advantages of the process over existing thermal drying and mechanical
dewatering systems.
6.2 Introduction
Essentially all coal supply agreements impose strict limitations on the amount of moisture
contained in the shipped product. Residual moisture lowers heating value, increases
transportation costs and can create downstream handling/freezing problems for customers. To
meet the moisture specification, a variety of solid-liquid separation processes are used in modern
coal preparation plants. Available methods for reducing surface moisture can be broadly
183 |
Virginia Tech | classified into three main groups: sedimentation, filtration and thermal drying (Wills and Napier-
Mun, 2006). Sedimentation methods make use of static or induced centrifugal forces to separate
solids from water based on differential settling/compaction, while filtration methods trap
particles against a mesh or porous medium to separate solids from water. Equipment such as
vibrating screening systems and various types of centrifugal dryers (stoker, screen-scroll and
vibratory centrifuges) are commonly used to dewater coarser coal particles. Finer coal particles
(<0.5-1 mm topsize) are typically dewatered using more complex dewatering equipment such as
screenbowl centrifuges and various types of vacuum disc and belt filters (Luttrell et al., 2007).
Unfortunately, existing fine coal dewatering processes are inefficient in terms of moisture
reduction, solids recovery and/or energy consumption (Osborne, 1988; Le Roux et al., 2005;
Keles, 2010).
It is widely recognized that the moisture content attainable by mechanical dewatering
systems is strongly dependent on coal particle size. For example, Figure 6.1 shows the
approximate lower limit on moisture than can be attained using mechanical coal dewatering
equipment.The inverse relationship between particle size and moisture content should be
expected due to the sharp increase in surface area as particle topsize is reduced. The finest coal
fraction can account for as little as a few percent by weight of the total run-of-mine coal, but may
represent one-third or more of the total moisture in the final coal product. In some industrial
operations, fine (<100-200 micron) or ultrafine (<40-50 micron) coal particles may be intentially
removed by classification circuits and discarded at the plant site to avoid an unacceptably high
product moisture. This loss represents a waste of valuable coal resources and a potential
environmental liability when discarded into waste impoundments (Orr, 2002).
184 |
Virginia Tech | 100
80
60
40
30
Vacuum/Pressure Filters
20
Screen-Bowl Centrifuges
Screen-Scroll
10 Centrifuges (Fine)
8
Vibratory Centrifuges
6
Screen-Scroll
4
Centrifuges (Coarse)
3
Vibratory Stoker
Historically Only Vibrating
Centrifuges
2 Attainable Using Screens
Thermal Dryers
1
325M200M 100M 48M 28M 14M 8M6M4M¼” ½” 1” 2” 4” 8”
Particle Size (Inches or Mesh)
Figure 6.1 Comparison of dewatering alternatives for different particle size ranges.
Historically, thermal dryers have been utilized in the coal prepration industry to reduce
clean coal moisture to single-digit values whenever mechanical dewatering processes were
incapable of meeting contract specifications. The most popular design is the fluidized bed dryer,
which uses coal, oil or gas as the fuel source to heat the incoming air stream. The amount of fuel
required depends on the amount of water that must be evaporated which, in turn, depends on the
amount of coal fed to the dryer and the percentage of water in the dewatered product (Miller,
1998). When operating correctly, thermal dryers can reduce the clean coal moisture to less than
6% by weight (Meenan, 2005). Unfortunately, thermal dryers involve a substantial investment of
upfront capital funds when installed and large annual costs for equipment maintenance and repair
throughout their lifespan. Operating costs for thermal dryers have also greatly increased in recent
years in response to higher fuel and labor costs. Thermal dryers can also suffer from emission
)%(
erutsioM
tcudorP
185 |
Virginia Tech | problems associated with fugitive dust and poor opacity. In fact, the opacity standard for coal
dryers was recently reduced from 20% to 10% as a result of a recent legislative action. Emissions
of nitrous oxides, sulfur dioxide, volatile organic compounds (VOCs) and particulate matter may
also present issues for some sites seeking operating permits. Moreover, thermal drying of
combustible particles of coal can present safety hazards resulting from accidental fires or dust
and gas explosions.
The development of an innovative, efficient and low cost technology for removing
moisture from fine coal is an important need for the coal preparation industry. In light of this
need, a novel non-thermal, mechanical dewatering process has been developed by NDT™ for the
coal preparation industry. In the current study, an experimental test program was undertaken to
evaluate the dewatering performance of the NDT™ process. This article provides a brief
description of the new patent-pending dewatering technology and presents experimental results
obtained from recent bench- and pilot-scale test programs.
6.3 Nano Drying Technology
The NDT™ drying system uses molecular sieves to wick water away from wet fine coal
particles and does not require crushing or additional finer sizing of the wet coal to dry it. These
molecular sieves are a form of nano-technology based particles, which are typically used for
extracting moisture from airborne, aerosol and liquid environments. There are also known
techniques for combining molecular sieves with solids, but no previous techniques included
regeneration of the molecular seives. Molecular sieves contain pores of a precise and uniform
size, typically in the range of 3 to 10 angstroms (Ramakrishna, Ma and Matsuura, 2011). These
pores are large enough to draw in and absorb water molecules, but small enough to prevent any
of the fine coal particles from entering the sieves. Some molecular sieves can adsorb up to 42%
186 |
Virginia Tech | of their weight in water (Bland et al., 2011). Molecular sieves are used in the drying process
because these are re-usable after the absorbed water is removed from the sieves by heating.
Molecular sieves often consist of alumino silicate minerals, clays, porous glasses, micro-
porous charcoals, zeolites, active carbon or synthetic compounds that have open structures
through or into which small molecules such as nitrogen and water can diffuse (Breck, 1964).
When the molecular sieves are mixed with wet coal fines, these sieves quickly draw water away
from the wet solids. In order to maximize surface contact between molecular sieves and coal
particles, the mixture is contacted/mixed/agitated for a short period of time. After contacting, the
molecular sieves are recovered from the dry coal by simple screening since the sieves are
substantially larger in size than the topsize of the dried coal particles. Once the separation occurs,
the remaining coal particles have a substantially reduced moisture content, which can reach low
single-digit values regardless of coal particle size. The molecular sieves are then regenerated by
removing the trapped moisture and are recycled back through the process. It is important to note
that the regeneration occurs after the deeply dewatered coal particles have been removed (i.e., no
portion of the coal is ever subjected to heating). Consequently, this process is considered by the
inventors to be an advanced dewatering process and not a thermal drying process, which offers
many advantages in terms of operational cost and environmental compliance.
6.4 Bench-Scale Testing
6.4.1 Experimental Procedure
A bench-scale experimental test program was performed to evaluate the performance of
the process of the NDT™ system in removing water from fine coal. For all experimental tests,
the wet feed sample consisted of either 0.6 mm or 0.15 mm topsize clean metallurgical coal
(filter cake) collected from an industrial plant.
187 |
Virginia Tech | Figure 6.2 Schematic diagram of the bench-scale NDT™ process.
During testing, a weighted sample of as-received fine feed coal was mixed with a
predetermined weight of molecular sieves. The mixture was then contacted together in a small
bench-scale rotary mixer for a defined period of time (Figure 6.2). After contacting, the mixture
of molecular sieves and coal fines was separated by using a laboratory sieve. The dewatered coal
particles passed through the sieve and were collected as an underflow product, whereas the
molecular sieves were retained on top of the sieve and were collected as an overflow product.
Once separated, the coal particles and molecular sieves were individually weighed and the
reduction in the percentage moisture of the coal sample was calculated. The last step in the
experimental procedure was drying the molecular sieves. To speed the regeneration process, a
microwave oven was used to evaporate the moisture held in the pores of the molecular sieves.
The regenerated molecular sieves were then reused in the testing program. No significant
188 |
Virginia Tech | Table 6.1 Overview of parametric tests conducted using the NDT™ process.
Experimental Experimental Particle Media Batches/Group Tests/Batch
Group Design Type Top-size Type (Size) (Number) (Number)
A Exploratory 0.6 mm I 5 8
B Central Composite 0.6 mm I 1 39
C Central Composite 0.6 mm I 1 39
D Uniform Grid 0.15 mm II 4 12
E Central Composite 0.15 mm II 1 52
difference was observed in the effectiveness of the moisture removal using either newly
manufactured or regenerated molecular sieves.
Five independent “groups” of statistically designed bench-scale experiments were
performed using the patent-pending process developed by NDT™ (Table 6.1). The type (size) of
molecular sieves and weight of coal sample was kept constant for each experimental group,
while the weight of molecular sieves and time of contact were varied over a range of
predetermined values as dictated by the statistical parametric test matrix. Duplicate test runs (a
minimum of 3 to 4) were conducted at each test point to assess the degree of variability and level
of reproducibility in the test data. The first group of tests (Group A) were comprised of
exploratory tests designed to identify the suitable ranges of experimental conditions for testing.
This group of test runs involved the processing of 5 batches of sample with 8 experimental test
runs per batch. Groups B and C consisted of two sets of central composite designs of 39 tests
each (15 central point tests). These groups were identical except for the range of variables
examined. Groups D and E were conducted using a different type (size) of molecular sieve. The
test matrix for Group D consisted of a uniform grid with 4 batches of experiments involving 12
test runs each, while Group E consisted of a central composite design encompassing a single
189 |
Virginia Tech | batch of 52 test runs (20 central point tests). After completing each test matrix, the data were
evaluated using standard statistical techniques.
6.4.2 Results and Discussion
A target moisture of 9% was selected with a range of 8 to 10% as the operating parameter
for the process of the NDT™ system . When 100 mesh x 0 product coal gets below 8% moisture,
dust problems become a concern and, if dried further, then explosion hazards must be
considered. If the moisture is more than 10%, then the potential benefits of adding this size
mateial to the clean coal are reduced. Therefore, the tests were designed to determine whether
the process of the NDT™ system could produce a 9% moisture product with a 95% confidence
level. The central points for Groups B, C and E were specifically selected and each central
composite design was statistically configured to see if this 95% confidence level could be
obtained for a 9% product moisture. It should be noted that maximum drying tests conducted
during the bench- and pilot-scale testing showed that moisture levels in the 1.5 to 2.5% range
could be easily produced if desired. Table 6.2 shows the overall performance of the NDT™
process in terms of average moisture contents of products for each batch/group. The data
indicates that the technology can readily provide single-digit moistures over a wide range of
operating conditions. In fact, moisture values in excess of 10% were only obtained when using
very short contact times or when low weight fractions of molecular sieves were utilized. To fully
demonstrate the impact of these factors, one group of tests from type I (i.e., Group B) and one
group of tests from type II (Group E) were selected for further discussion in this publication.
Figure 3 shows the central composite text matrix used in the Group B test programon the 0.6 mm
x 0 feed
190 |
Virginia Tech | used in the 15 central point tests was 21.8+0.16% with a standard deviation of 0.90. After
contacting with the molecular sieves, the product moisture dropped to a average value of
8.90+0.02% with a standard deviation of 0.14. The very small confidence interval and low
standard deviation values associated with the data obtained for the dewatered product indicates
that a high degree of reproducibility can be achieved using the bench-scale version of the process
of the NDT™ system.
A similar trend in moisture removal was observed for the tests conducted for Group E
having a topsize of 0.15 mm. These experiments were conducted using type II molecular sieves
over a similar range of contact times and a lower range of media factors. Each satellite test
conducted around the central test point was repeated 4 times to assist in identifying outliers and
evaluating reproducibility. The central test point, which involved a contact time of 3.5 minutes
and media factor of -0.63, was repeated 20 times in random order throughout the test matrix. For
this particular group of tests, the average moisture contents of the as-received 0.15 mm x 0 feed
was 26.2+0.10%. After contacting with the molecular sieves, the 0.15 mm x 0 product moistures
were reduced to single-digit values for all tests conducted at contact times of 3.5 minutes or
longer (see Figure 6.5). The lowest product moisture content of 6.38% was achieved for the
longest contact time of 4.9 minutes. Tests conducted with contact times less than 3.5 minutes did
not achieve single-digit mositures, but at 10.2-10.9% moisture were not far from breaking this
meaningful barrier.
One noteworthy difference in the Group E test series was the greater degree of scatter in
the experimental data. Standard deviation values greater than 1 were observed for the vast
majority of the test points and a value as high as 4.67 was obtained for one of the satellite tests.
193 |
Virginia Tech | 6.5 Pilot-Scale Demonstration
In light of promising bench-scale data, a decision was made to construct a pilot-scale
NDT™ plant to demonstrate the capabilities of this new patent-pending technology in continous
mode. While the small scale testing validated the basic system, numerous additional proprietary
refinements were developed by NDT™ for operating on a larger scale. The flowsheet for the
facility is shown in Figure 6.7.
Feed Coal/Sieve
Coal Contactor
Make-Up Coal/Sieve Dry
Sieves Screen Coal
Water Sieve
Vapor Regenerator
Figure 6.7 Simplified flowsheet for the pilot-scale NDT™ processing facility.
The completed facility, which was largely assembled using off-the-shelf components,
was designed with an effective throughput capacity of 1,000 pounds per hour (0.5 TPH). The
self-contained facility included unit operations for handling, contacting and separating the coal
and media. An advanced gas-fired dryer was used to regenerate the molecular sieves such that
the entire process operated in a closed-circuit loop. The prototype facility was designed,
constructed and successfully commissioned over a period of approximately 10 months. During
this time, shakedown tests were completed and the process circuit was refined, modifed and
optimized using proprietary optimization techniques to provide a demonstration facility that
operated smoothly and efficiently.
195 |
Virginia Tech | Table 6.3 provides an overview of test results obtained with various coals using the pilot-
scale NDT™ facility. As shown, the prototype facility successfully achieved single-digit product
moistures for a wide range of feed coal applications. Engineering criteria developed from bench-
scale testing, such as contacting (retention) times and coal-to-sieve loadings, were also validated
using the pilot-scale plant. More importantly, the pilot-scale test runs successfully demonstrated
that the molecular sieves could be regenerated and recyled back through the process without
incuring significant losses due to media degredation and at a lower heating/evaporation cost than
traditional thermal drying.
Table 6.3 Examples of pilot-scale NDT™ test results.
Coal Particle Capacity Feed Product
Source Size Class (lb/hr) Moisture (%) Moisture (%)
A 1 mm 1,600 17.88 7.63
B 1 mm 1,200 10.41 5.38
1 mm 1,200 10.41 7.13
1 mm 1,200 10.41 6.84
C 0.15 mm 600 27.28 2.52
0.15 mm 550 27.28 7.46
D 0.15 mm 1,000 31.83 3.18
0.15 mm 1,000 31.83 5.86
0.15 mm 1,000 31.83 8.27
6.6 Discussion
The removal of unwanted moisture from fine coal has historically been considered one of
the most challenging technical problems in the coal preparation industry. The process of the
NDT™ system was developed specifically to address this issue by providing (i) effective
moisture removals, (ii) efficient energy utilization and (iii) enhanced environmental
196 |
Virginia Tech | Recent estimates by an environmental consulting group indicate that emission reductions
as large as 90% or more when compared to a thermal dryer are possible using the NDT™
system. The emission projections from one such case study is shown in Figure 6.8. In this case,
emissions of volatile matter (VM), sulfur dioxide (SO2) and particulate matter (PM) would be
essentially elimiated (>99% reduction) using the NDT process. Projected emissions of carbon
dioxide (CO2) and nitrous oxides (NOx) would be reduced by 91% and 84%, respectively. It is
particularly important to note that the projected total emissions of 59.4 tons per annum (TPA) of
criteria pollutants is likely to be less than the threshold value that would trigger the need for a
Title V Air Quality permit in many states. For example, no such permit would be required in
West Virginia since the threshold value is 100 TPA of criteria pollutants. The process of the
NDT™ system also generates no other by-products that could potentially be released into the
environment.
Finally, it should be noted that the NDT™ drying system is very efficient in terms of
energy utilization. Since only the molecular sieves are dried, the drying step can be fully
optimized in the absence of coal-imposed contraints associated with dryer temperature levels,
gas-solid contacting systems, and coal dust explosions. As such, the system provides the highest
possible energy efficiency at the lowest possible fuel cost. Since the process treats only the fine
coal fraction, which is generally between 10 to 15% of the total clean coal product (and not the
entire clean coal product treated by conventional thermal dryers), the required footprint for the
facility is only a fraction of that demanded by a large-scale coal thermal dryer. Also, due to
fewer operational complexities, significant cost savings are also expected for ancillary items
such as electricity, chemicals, maintenance and labor. Cost estimates conducted in cooperation
with a commercial engineering firm are plotted in Figure 6.9. Although such economic
198 |
Virginia Tech | calculations tend to be site specific, the costing figures for this site do suggest a relative
operating cost of less than half of that required to operate a conventional thermal dryer.
6.7 Conclusions
The removal of surface moisture from fine coal has been a longstanding problem in the
coal industry. To address this need, an innovative process based on nano-technology has been
developed. Bench-scale studies indicate that the Nano Drying Technology (NDT™) proprietary
system provides an effective method for coal drying. The NDT™ system can effectively
dewater fine (1 mm x 0) coal from slightly more than 30% surface moisture to single-digit
values. Test data obtained using a pilot-scale NDT™ plant further validated this impressive
capability using a continous prototoype facility. It was also observed that, unlike existing fine
coal dewatering processes, the performance of the NDT™ system is not dictated or constrained
by particle size, i.e., it works equally well on 1 mm x 0 coal as it does on 325 mesh x 0 coal. The
process of the NDT™ system overcomes problems associated with other techniques for fine
coal drying since dewatering occurs at ambinent temperature and low airflow. Only the
molecular sieves have to be dried, which reduces energy. Moreover, this process produces no
damaging contaminants and has a very small installed footprint and environmental impact.
199 |
Virginia Tech | CHAPTER 7 SUMMARY AND CONCLUSIONS
To improve the separation efficiency of fine coal cleaning circuits, several series of
laboratory-scale, pilot-scale and field tests were conducted using different fine coal cleaning
technologies/circuits. Based on the results obtained from this work, engineering criteria based on
feed size characteristics and sulfur contents was developed to identify optimum circuit
configurations for the processing different fine coal streams.
In the first phase of work, different laboratory and pilot scale test circuits (using spirals,
water only cyclones, teeter bed separators, hydroFloat and froth flotation) were constructed for
the purpose of conducting a detailed experimental study on the separation efficiency of fine coal
cleaning processes. The results obtained from the study were then used to identify optimum coal
particle size ranges for maximum separation efficiencies for different fine coal cleaning
technologies. The data obtained from this work indicates that the most effective processes for
each size range were generally (i) froth flotation for feeds finer than about 0.3 mm, (ii) spirals for
feeds sized to 1 x 0.3 mm, and (iii) teeter-bed systems (particularly the HydroFloat™
technology) for feeds larger than 1 mm. Water-only cyclones were not found to be effective as
stand-alone units due to the potential for high coal losses when secondary back-up units are not
available within the plant circuitry.
In the second phase of work, pilot-scale and in-plant testing was conducted to identify
new types of spiral circuit configurations that improve fine coal separations. Five different spiral
circuits were constructed and experimentally tested at the pilot-scale to evaluate their separation
performances. The experimental data thus obtained indicates that a four-stage spiral with second-
201 |
Virginia Tech | and fourth-stage middlings recycle offered the best option for improved separation efficiency,
clean coal yield and combustible recovery. The newly developed spiral circuitry was capable of
increasing cumulative clean coal yield by 1.9 % at the same clean coal ash as compared to that
achieved using existing conventional compound spiral technology. Repluping of coal slurry after
two turns proved to be ineffective in improving the separation performance of spiral circuits.
In the third phase of work, various methods were investigated for improving the rejection
of both ash-bearing minerals and sulfur-bearing pyrite from fine coal cleaning circuits. The
experimental findings from both laboratory and pilot-scale tests indicated that density-based
separations are generally effective in reducing fine coal sulfur due to the large density difference
between pyrite and coal. The data also showed that sulfur rejections obtained in flotation-only
circuits were often poor due to the natural floatability of pyrite. Unfortunately, engineering
analyses showed that pyrite removal from the flotation feed using density separators would be
impractical due to the large volumetric flow of slurry that would need to be treated. On the other
hand, further analyses indicated that the preferential partitioning of pyrite to the underflow
streams of classifying cyclones and fine wire sieves could be exploited to concentrate pyrite into
low-volume secondary streams that could be treated in a cost effective manner to remove pyrite
prior to flotation. Therefore, on the basis of results obtained from this experimental study, a
combined sieve screen-flotation-spiral circuitry was developed for enhanced ash and sulfur
rejections from fine coal circuits.
In the fourth and final phase of work, experimental tests were carried out to investigate a
new mechanical, non-thermal dewatering process called Nano Drying Technology (NDT™).
Results obtained from bench-scale testing showed that the NDT™ system can effectively
dewater fine (1 x 0 mm) clean coal products from more than 30% surface moisture to single-digit
202 |
Virginia Tech | FORCE AND ENERGY MEASUREMENT OF BUBBLE-
PARTICLE DETACHMENT
by
Hubert C.R. Schimann
R.H. Yoon, Chairman
Mining and Minerals Engineering
(ABSTRACT)
Possibilities for increasing the upper limit of floatable particle sizes in the froth flotation
process have been examined since the early beginnings of mineral flotation. The
economic implications of such an increase are far ranging; from decreased grinding costs
and increased recoveries to simplified flow-sheet design and increased throughput, all
leading to increased revenue. Bubble-particle detachment has been identified as the main
limiting factor for coarse particle flotation. The detachment process has been studied to
better understand the factors influencing the strength of attachment and the energies
involved. Direct measurements of bubble particle detachment were performed using a
hanging balance apparatus (KSV Sigma 70 tensiometer) and using a submerged
hydrophobic plate in water. Three experiments were used; direct force measurement of
bubble-particle detachment, detachment force and energy of a bubble from a submerged
hydrophobic plate, and detachment force and energy of a cetyltrimethylammonium
bromide coated silica sphere from a flat bubble. Octadecyltrichlorosilane was used
as a hydrophobic coating in the first two experimental methods. These experiments were
recorded with a CCD camera to identify the detachment processes involved. Energies for
both methods were calculated and divided into the two main steps of the detachment
process: Three-Phase-Contact pinning and three phase contact line sliding. The first step
represents the energy barrier which must be overcome before detachment can begin. It is
directly related to contact angle hysteresis. Detachment occurs during the second step,
where the solid-vapor interface is replaced by solid-liquid and liquid-vapor. This step |
Virginia Tech | ACKNOWLDEGEMENTS
First and foremost, the utmost appreciation is extended to my advisor Dr. Roe-Hoan
Yoon, for his guidance, inspiration and support throughout the course of my
investigations. The independence and freedom in choosing the direction to follow with
my projects was also much appreciated. Special thanks are also extended to Dr. Jan
Christer Eriksson (Department of surface chemistry, Royal Institute of Technology,
Stockholm, Sweden) for the many conversations and advice which were instrumental in
developing the ideas presented here. I am also grateful to Dr. Tom Novak, Dr. Gerald
Luttrell, and Dr. Greg Adel who as members of my committee provided pertinent
feedback. Special appreciation is given to Dr. Jinming Zhang, Jin Hong Zhang, and
Jialin Wang for their helpful discussions, suggestions, friendship and willing assistance.
Thanks are in order to Mariano Velázquez for his technical assistance and friendship.
I would also like to thank Dr. Dimetri Telionis and Dr. William Ducker for the courteous
loan of equipment and lab space.
Particular thanks are in order to my lab mates Mr. Emílio Lobato, Mr. Ian Sherrell and
Mr. Selahattin Baris Yazgan for all their helpful suggestions and continued support
through all my ups and downs. I am most grateful for their friendships that were built
over the past two years.
I owe special gratitude to Mr. Mert Kerem Eraydin for his friendship, continuous support
and encouragements throughout my stay at Virginia Tech.
I would like to express my most sincere appreciation to my parents Monique and Karl
Schimann for their inspiration, encouragements, wisdom, and continued suggestions. I
would also like to thank my sister Erika and brother Michaël for their continued support
throughout my two years here.
iv |
Virginia Tech | INTRODUCTION
Coarse particle flotation
Froth flotation is widely used in the mining industry to separate valuable minerals from
other materials in their host environment. Minerals are separated by attaching themselves
to rising air bubbles in the flotation cell and then recovered at the top of the cell. The ore
must be ground small enough so that flotation can proceed (e.g. <.5 mm diameter).
Coarse particle flotation provides opportunities for reduced grinding costs, increased
recoveries and simplified flow-sheet designs (by eliminating certain classifying and other
steps); all leading to increased throughput. However, the limits of coarse particle
flotation have been well demonstrated as a severe decrease in recovery above a certain
size with plant and laboratory batch tests data (Trahar 1981). A single bubble in a
Hallimond tube was also used to demonstrate the dramatic drop in recovery with
increasing particle size (Drzymala 1994). While other work identified a flotation domain
from the minimum contact angle at which individual particle sizes will float (Crawford
and Ralston 1988).
Flotation is a function of the probability of particle collection (Sutherland 1948)
P = P ⋅P ( 1−P ) [1]
C A D
where P , P , and P are the probabilities of collision, adhesion, and detachment
C A D
respectively. The probability of collision depends on hydrodynamic effects in the
flotation cell. The probabilities of adhesion and detachment are a combination of
hydrodynamic effects and surface chemistry of the bubble and particle. Since surface
chemistry is not affected by particle size, the probability of adhesion does not change
much with particle size. Probability of detachment is then the main limiting factor in
coarse particle flotation.
1 |
Virginia Tech | Research has focused on characterizing the various factors in flotation in attempts to
increase the maximum floatable particle size. Early work identified the work of adhesion
W =T ( 1−cosθ) [2]
wa
as the work done per unit area to create the air-solid interface at the expense of the solid-
water interface (Wark 1933). Where T is the surface tension of water and θ is the
wa
contact angle between the water and solid. Wark (1933) also attempted with limited
success to calculate the maximum floatable particle size. The force and work required to
remove a particle from the liquid-vapor interface was later measured using a centrifugal
force apparatus and calculated as a function of the solid and liquid densities (Nutt 1960).
Mathematical models of the flotation process followed taking into account the various
sub-processes and physico-chemical properties involved in the system (Mika and
Fuerstenau 1968). However, this model still suffered from a lack of proper description of
these sub-processes to properly describe flotation. A theoretical evaluations of the upper
particle size in flotation was calculated (Schulze 1977). These calculations were used to
produce a rough estimation of the energy of rupture of a particle from a bubble.
However, the work calculated in this manner does not represent an accurate description
of the energy involved in the detachment process. A theoretical detachment force model
was developed but disagreed by one order of magnitude when compared to forces
measured using the centrifuge method (Nishkov and Pugh 1989).
More recent research on the stability of the bubble-particle aggregate in a flotation
column using laboratory batch tests established some qualitative parameters for
detachment energy of a particle from an oscillating bubble (Falutsu 1994). A bubble
vibration detachment force measurement technique was developed to more closely
resemble the conditions in flotation (Cheng and Holtham 1995), which compared
favorably with previous theoretical work (Nutt 1960). Attachment and detachment
efficiencies have also been described from calculation of the bubble-particle interaction
forces in flotation and used to identify a maximum floatable particle size (Ralston et al.
1999).
2 |
Virginia Tech | Turbulent forces in the cell were recently modeled as the cause for bubble-particle
detachment (Bloom and Heindel 2002; Pyke et al. 2003). Similarly, the detachment at
the pulp-froth interface was described from experimental data by proposing an empirical
model of percent apparent detachment during flotation (van Deventer et al. 2004). This
later model compared attached and recovered particles to estimate the detachment.
Detachment force measurements
Direct measurements of detachment forces have been studied to better understand the
detachment process and its implications for coarse particle flotation. Understanding of
the detachment process holds the key to increasing the upper size limit of floatable
particles. The detachment force of a particle from a bubble has been measured to
correlate it with contact angle, and interfacial energy values (Janczuk 1983; Janczuk
1985; Janczuk et al. 1990). However, these all use controlled contact areas which
prevent the bubble from freely spreading on the particle surface. This may represent
some cases encountered in flotation but fails to accurately describe the overall interaction
between bubbles and particles. A centrifuge method has also been used to measure the
detachment force (Schulze et al. 1989). However, this method does not allow for
observation of the detachment process, providing only a force range at which particle
detachment occurred.
The atomic force microscope has provided an ideal method for measuring interactions
between bubbles and individual colloidal particles(Butt 1994; Butt et al. 1995; Preuss and
Butt 1998a; Preuss and Butt 1999). However, they have proved more useful for
measuring interactions during the approach cycle than detachment. The amount of
bubble deflection during detachment is difficult to accurately measure, thus the
detachment cycle cannot by accurately described using this method.
The detachment models proposed above describe the detachment force with varying
success, but do not accurately provide the detachment energy process. In the work
3 |
Virginia Tech | Direct force and energy measurement of
bubble-particle detachment
Hubert C.R. Schimann
Department of Mining and Minerals Engineering, Virginia Polytechnic Institute and State University,
Blacksburg, VA, 24061, USA
Abstract
Bubble-particle detachment is the main limiting factor in coarse particle flotation.
Detachment occurs when forces on the bubble-particle aggregate in a flotation cell
overcome the strength of adhesion. Strength of adhesion depends on the particle surface
roughness, surface chemistry, and surface tension of the liquid media. In the present
work, detachment forces between spherical particles of varying hydrophobicity and air
bubbles were directly measured using a modified KSV 70 surface tensiometer. The
experiment was recorded with a 1 kHz CCD camera to monitor changes in bubble-
particle contact area, bubble size and shape, and “neck formation”. The detachment
energy was found to consist of three parts; bubble stretching, bubble sliding, and bubble
necking. The measured energies compared favorably to the work of adhesion.
Keywords: detachment, bubble – particle interaction, coarse – particle flotation,
octadecyltrichlorosilane
Introduction
Froth flotation is widely used in the mining industry to separate valuable minerals from
other materials in their host environment. Minerals are separated by attaching themselves
to rising air bubbles in the flotation cell and then recovered at the top of the cell. The ore
must be ground small enough so that flotation can proceed (e.g. <.5 mm diameter).
Coarse particle flotation provides opportunities for reduced grinding costs, increased
recoveries and simplified flow-sheet designs (by eliminating certain classifying and other
steps); all leading to increased throughput.
Flotation is a function of the probability of particle collection (Sutherland 1948);
10 |
Virginia Tech | P = P ⋅P ( 1−P ) [3]
C A D
where P , P , and P are the probabilities of collision, adhesion, and detachment
C A D
respectively. The probability of collision depends on hydrodynamic effects in the
flotation cell. The probabilities of adhesion and detachment are a combination of
hydrodynamic effects and surface chemistry of the bubble and particle. Since surface
chemistry is not affected by particle size, the probability of adhesion does not change
much with particle size. Probability of detachment is then the main limiting factor in
coarse particle flotation.
Theory
Flotation efficiency can be described largely in terms of contact angle or particle
hydrophobicity. The equilibrium contact angle, θ , is the angle formed between a
Eq
bubble and solid surface in water once the bubble becomes stable. It is described
thermodynamically by Young’s equation.
γ −γ
SV SL =cosθ [4]
γ Eq
LV
Where, γ , γ , and γ represent the solid-vapor, solid-liquid, and liquid-vapor
SV SL LV
interfacial tensions respectively. Where the liquid, vapor, and solid phases meet is
referred to as the three-phase-contact (TPC) point or line. The interaction between two
materials 1 (air) and 2 (solid) immersed in a third liquid 3 is given by (van Oss 1994);
∆G =γ −γ −γ [5]
SV SL LV
11 |
Virginia Tech | Combining equations [4] and [5] leads to the Young-Dupré equation, which gives the free
energy change for the attachment of a bubble onto a solid surface.
−∆G =γ ( 1−cosθ) [6]
LV
This represents the work required to remove liquid from the solid surface and bubble
surface and create a new solid-vapor interface. This change in free energy is referred to
as the work of adhesion since it represents the work required to form a bubble-particle
aggregate. Calculating the actual energy spent requires the addition of an area term to
equation [6] to account for the various initial and final interfacial areas. The
thermodynamics of the bubble-article aggregate formation are reversible so that work of
adhesion can be used to describe the energy required for detachment as well.
Bubble-particle detachment in flotation is mostly caused by turbulence in the cell. Once
the forces on the particle are larger than the adhesion force, the particle detaches.
The detachment process depends on contact angle, media surface tension and surface
heterogeneity or contact angle hysteresis. Contact angle determines the contact area
between bubble and particle, which sets the distance that the TPC line will have to travel
on the surface before detachment occurs. Surface tension corresponds to the tensile
strength of the bubble.
Figure 1 - Advancing and receding contact angle
Contact angle hysteresis has a role in determining the force and energy that must be
overcome before the detachment process can begin. Hysteresis is caused by surface
12 |
Virginia Tech | roughness (Israelachvili 1991) and on a smaller scale by surface energy of the solid
(Chibowski 2003). When a liquid drop moves on a surface, the angle formed at the side
of the drop which is advancing on the surface is referred to as the advancing contact
angle (see Figure 1). And reversely, the angle formed at the receding end is the receding
contact angle. If a bubble is detached from a particle, the TPC will not move until the
advancing contact angle is reached. At the higher contact angle the water can advance on
the solid surface. The energy spent shifting the TPC from equilibrium to advancing angle
represents energy barrier which must be overcome before the detachment process can
begin. Alternatively, hysteresis also raises some questions about the reversibility of the
work of adhesion. It is believed that during the recession process, the interface may not
be retracing its original path, so that the process may not be thermodynamically
reversible (Israelachvili 1991). This will be revisited in the results and discussion.
Surface tension is an important determining factor in bubble-particle adhesion and
detachment. It determines the strength and elasticity characteristics of the bubble. A
lower tension means a less ‘stretchable’ but stronger bubble. A more ‘stretchable’ bubble
may require more energy, but less force to detach.
Experimental Methods
Forces were measured using a Sigma 70 surface tensiometer (KSV Instruments Ltd.).
This equipment was a hanging balance with a resolution of 1µN. The bubble-particle
interactions were recorded using a Phantom V4.0 CCD camera (Photo-Sonics Inc.).
Interactions were also photographed by a 4.0M pixel S4 digital camera (Canon Inc.)
equipped with a reversed 50mm AF NIKKOR (Nikon Corporation) lens (which allows it
to act as a macro lens).
All the reagents used in the experiment were at least ACS grade and were obtained from
either Fisher Scientific or Alfa Aesar. Soda lime glass spheres with a mean diameter of
2007µm ±40µm (Duke Scientific Corporation) were methylated with
13 |
Virginia Tech | octadecyltrichlorosilane (OTS) and butyltrichlorosilane (BTS) to create surfaces with
varying hydrophobicity. The methylation procedure consists of washing the samples for
1hr in Piranha solution (30% H O :70% H SO ) at 60-70°C. The samples are then rinsed
2 2 2 4
in Nanopure water and immersed in a 10-5M silane solution in toluene for 30-90 minutes
depending on the target contact angle. The samples are then removed from solution and
first rinsed with chloroform to remove any excess toluene from the surface, then acetone
to remove any physisorbed silane. Following this, the samples are placed in acetone in
an ultrasonic bath for at least 20min. This breaks up any amalgams of polymerized silane
which may have accumulated at the silica surface, thereby creating a more uniform silane
coating. The samples are stored in Nanopure water in sealed containers until the
experiment.
A sphere was fixed to the end of a glass hook, which was manufactured by the Virginia
Tech glass shop, using Crystalbond™ 509 glue (Electron Microscopy Sciences). The
sphere was then suspended from the tensiometer above the bubble, which sat on the end
of a thick glass rod (see Figure 2). The bubble support sat in a glass cell filled with water
on top of a mechanical stage that could be moved vertically at 1mm/min to bring the
bubble and particle into contact and then detach them.
14 |
Virginia Tech | Figure 3 is a picture of the air bubble attached to its support as it is being pulled away
from a glass sphere.
Figure 4 - Detachment force curve Figure 5 - Round bubble detachment
force
A typical output curve from the surface tensiometer is shown in Figure 4. The
detachment force is measured as the difference between the maximum and the baseline.
The baseline is the force at which the sphere is detached. This is the force exerted on the
instrument from the weight of the sphere only. The energy is taken as the integral of the
force applied across the distance from equilibrium to detachment. Equilibrium is shown
in the figure as x . It is the point at which the bubble is neither pushing up nor pulling
0
down on the sphere.
Detachment force measurement between the sphere and bubble were not very
reproducible when using a bubble of similar size as the sphere. Figure 5 displays force
measurements performed under these conditions at varying contact angles.
In the next series of measurements, the small bubble was replaced by an infinitely large
(or flat) bubble. The flat bubble removes the variations in bubble radius. The air bubble
was created using a PTFE rod, with a cone bored out in the centre, connected to a syringe
that provided the air for the bubble, see Figure 6.
16 |
Virginia Tech | Figure 8 - Detachment force and contact area of
methylated sphere from flat bubble (contact angle = 95°)
In the first component, the three-phase-contact (TPC) line is pinned on the silica surface
while the bubble is being stretched. The bubble stretches until the advancing contact
angle is reached. Surface roughness may cause the bubble to stretch slightly beyond this
point. Once the advancing contact angle is reached, the vapor phase retracts from the
sphere and the bubble commences detachment. The maximum force reached is
dependent on the magnitude of the hysteresis and on surface tension of the media. Figure
8 demonstrates the two stages of the detachment process. Contact area remains constant
(E ) until enough force is applied to move the TPC line and the advancing contact angle
1
has been reached. Once this occurs the force remains constant as the TPC line moves
‘down’ the sphere until detachment occurs.
The second component of the detachment process is dependent on particle
hydrophobicity because this determines the starting contact area between solid and vapor.
It also establishes a maximum, the advancing contact angle, which must be reached
before movement of the TPC line can begin. Comparison of the different energy
components with work of adhesion calculations further demonstrated this relationship.
18 |
Virginia Tech | Table 1 – Sphere-sphere energy of detachment and work of adhesion comparison
θ 48° 95°
E 1.824 x 10-2 J/m² 1.971 x 10-2 J/m²
1
E 1.794 x 10-2 J/m² 4.855 x 10-2 J/m²
2
E 3.618 x 10-2 J/m² 6.826 x 10-2 J/m²
total
W 1.367 x 10-2 J/m² 4.354 x 10-2 J/m²
a
Work of adhesion in the Table 1 was calculated from the differences in surfaces areas of
the various interfaces between the initial attached and final detached states.
Figure 9 - Sphere-sphere work of adhesion calculation
The relationship is solved by assuming the vapor volume does not change from the initial
to the final state. Solving the relationship gives the following equation for work of
adhesion.
∆A γ + A (γ −γ )
W = LV LV C SL SV [7]
A A
C
As can be seen from Table 1, the energy E , is fairly constant for 48° and 95° surfaces,
1
because this portion of the detachment does not depend on the surface chemistry
properties of the sphere. In mineral processing, this part of the detachment energy could
be maximized using frother to create stronger bubbles. E can also be regarded as an
1
activation energy for detachment due to surface roughness and the magnitude dependent
on the surface tension of the bulk solution in which the process takes place. E shows
2
19 |
Virginia Tech | good correlation with the work of adhesion calculation because it is the actual energy of
detachment whereas E is the energy necessary to begin the detachment process.
1
Detachment Mechanism Model
Sphere and flat bubble detachment measurements provided a better insight into the
detachment process, as bubble deformation was much more pronounced with the ‘large’
radius bubble. Figure 10 describes the bubble deformation geometry as the sphere is
pulled away.
R
(cid:72)
r (cid:137)
h tpc (cid:73)
Liquid
(cid:81)
Vapour
f
y
(cid:78)
Figure 10 - Flat bubble - sphere detachment geometry
The air-water interface deformation as the sphere is pulled away from the bubble can be
described by the following solution (Huh and Scriven 1969) to the Young-Laplace
equation for capillarity.
d2y dy2ρg dy2 1 dy
= 1+
y 1+
− [8]
dx2 dx
γ dx x dx
where ρ is the density of the liquid, g is gravity, and γ is the interfacial tension.
Equation [8] can be solved numerically and leads to the following solution describing the
height, h, of the TPC line above the flat bubble (James 1974).
20 |
Virginia Tech | 4L
h= Rsinαsinβ ln Rsinα( 1+cosβ)−σ
[9]
γ
L≡
ρg
R is the radius of the sphere, L is the capillary length and σ is the Euler constant, 0.577.
The force exerted by the bubble onto the sphere depends on the contact angle and shape
of the bubble as these determine the angle at which the force is applied. The total force F
exerted on the sphere in the vertical direction is
π
F =2πRsinα⋅γ cos −θ+α [10]
LV 2
When the sphere is at equilibrium position the bubble is not deformed and is not exerting
any force on the sphere. The angle between the sphere and bubble at this point is the
receding contact angle (Preuss and Butt 1998b), which can easily be calculated from the
contact radius. As the sphere is pulled away, the shape of the bubble is defined by
equation [8]. The contact angle is then calculated by solving equation [9] at progressing
heights of the sphere above the bubble. Once the contact angle is known, and assuming a
constant contact radius during the stretching portion of the detachment process, the force
is calculated using equation [10] for each sphere height above the flat bubble datum.
Force distance curves obtained in this manner were compared to the measured force
curves, see Figure 11.
21 |
Virginia Tech | Figure 11 – Detachment force VS sphere center distance from the flat bubble for (a)
38° (b) 78° (c) 95° methylated silica spheres
‘Calculated Force I’ represents the stretching portion of the force curve calculated as
described above. It represents the force exerted by the bubble on the sphere as the
contact angle shifts from receding to advancing contact angle. ‘Calculated Force I’
shows good correlation with the measured force curve, although, graph (a) demonstrates
some of the difficulty in simulating the detachment process. The experimental and
calculated forces are in agreement for the first third of the curve, at which point the
bubble seems to have ‘slipped’ on the sphere, causing a change in the slope due to the
new contact radius.
Once the advancing contact angle is reached the bubble then begins sliding off the
particle. This portion of the detachment curve was modeled by again solving equation
[9] at progressing sphere heights, but now assuming that θ stays constant (note that θ = α
+ β). This portion of the detachment process is represented by ‘Calculated Force II’,
which displayed good correlation with the measured force. Variation in this portion of
the curve is probably due to the assumption that θ is constant. This is not entirely correct
as the bubble will still experience some localized pinning on the sphere surface as it
slides off. The calculated force also assumes a clean break instead of the necking
behavior of the bubble seen at higher contact angles. Necking was observed during the
experiment as contact angle increased, as evidenced by the small bubble often remaining
22 |
Virginia Tech | on the sphere after detachment, see Figure 12. This behavior is clearly visible when
comparing graphs (a) and (c). In graph (a) there is almost no necking, so the measured
force drops off abruptly at the end of the curve. The force curve in graph (c) slowly
decreases before finally falling to zero.
SPHERE
BUBBLE
Figure 12 - Bubble on sphere surface after detachment
The measured energy was compared to the calculated work of adhesion for a sphere and
flat bubble interaction. Work of adhesion was calculated with the following equation
( )
W =γ πR2 sin2θ −2cosθ 1−cosθ [11]
A LV R Eq R
where R is the radius of the sphere and θ and θ are the receding and equilibrium
R Eq
contact angles respectively. The work of adhesion calculation describes the change in
free energies between the initial and final states as described in Figure 13.
Figure 13 - Sphere - flat bubble work of adhesion calculation
Equation [11] assumes that the initial state is receding contact angle, and the equilibrium
contact angle is used to cancel out the solid-liquid and solid-vapor interfacial tension
23 |
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