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Virginia Tech | Through correlation analysis, the underflow flow rate, overflow flow rate, and inlet
pressure are significantly dependent on the four input variables (Table 4.2). As expected, the
underflow opening has a positive effect on the underflow flow rate but negative impacts on the
overflow flow rate and inlet pressure. These relationships confirm the underflow’s function as a
restriction point similar to the ASH unit and typical hydrocyclone operation. The pedestal
dimensions have both positive and negative effects on the underflow and overflow streams
which indicate the pedestal as another source of restriction and can control the characteristics of
the product streams. The system feed rate is typical of hydrocyclone operation as increasing feed
rate has a positive effect on all three measured results. These correlations result in the optimum
design conditions that will achieve the desired overflow stream where minimal water reported.
Observed during the evaluation was the water level within the flotation cyclone, or better
described as the development of the air core. Drawing a conclusion about the air core using the
Design Expert software cannot be made due its non-linear behavior. Based on observations and
the results provided, an air core would develop twice while adjusting the input parameters. An
air core would first develop at lower feed rates or flow restrictions where the centrifugal forces
and water reporting to the overflow were minimal. However, as the feed rate or flow restrictions
increased, there would be an increase in water reporting to the overflow and the high centrifugal
Table 4.2. The Design Expert Correlation of System Inputs and Outputs.
Correlation of Inputs and Outputs
Correlation
Inputs Q (GPM) Q (GPM) ΔP (psi)
u o
D (in) 0.916 -0.808 -0.662
uf
D (in) 0.709 -0.669 -0.443
p
H (%) 0.11 -0.038 0.095
p
Q (GPM) 0.359 0.352 0.663
s
55 |
Virginia Tech | forces within the flotation cyclone would create another air core. In the case of centrifugal
flotation, the air core at lower feed rates and restrictions is desirable over the high centrifugal
field air core. These initial efforts of testing the secondary flotation cyclone design parameters
provided knowledge about the fluid mechanics and air core development which aids in the
design optimization for future coal testing.
Although the conclusions from the initial water evaluations of the flotation cyclone with
adjustable pedestal and axial underflow identified the design parameters as critical, not all of the
parameters could be considered adjustable during operation. Of the four parameters tested, feed
rate and underflow diameter could be considered the two adjustable operating variables.
Therefore, further testing to analyze the effects of varying feed rate and underflow diameter was
performed. Table 4.3 shows the results from the water evaluation of the feed rate and underflow
diameter. No air or frother was added to this evaluation to simplify the analysis.
Inspecting the results from the feed rate and underflow diameter test, similar conclusions
can be made about the correlation with product streams, inlet pressure, and water height. Table
4.4 shows the correlation coefficients relating the feed rate or underflow diameter to the
measured result. At the various underflow diameters, feed rate positively affects the product
streams and the inlet pressure. In comparison, by varying the underflow diameters, the underflow
flow rate increases as expected, which decreases the overflow flow rate and inlet pressure due to
the lower flow restriction. These conclusions result in the determination that underflow diameter
and system feed rate will be the main parameters to change in order to optimize the flotation
cyclone for coal evaluation. Once again in this designed experiment, the water level was not a
linear trend as two air cores could be developed due to low flow restriction and high centrifugal
forces.
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Virginia Tech | Concluding from the preceding tests, the underflow diameter provided a restriction point
that aided the development of the required air core for the flotation process and dictates the
overflow flow rate. Therefore, the optimum diameter could be found to set the desired air core
size and limit the amount of water in the overflow. However, observations during the initial
testing established that high inlet feed rates were required to maintain desired air core sizes and
overflow quantities. As stated in the design description, two cyclone lengths were constructed,
24 and 15 inches, for the evaluation of the secondary design. The objective of constructing the
shorter flotation cyclone was to aid the development of the air core by limiting the required feed
rate (Table 4.5).
The analysis of the 15-inch flotation cyclone shows the low pressure requirements and
the minimal amount of water reporting to the overflow stream despite the presence of air
addition. These results are desirable in a flotation process. The shorter cyclone acts in the same
manner where underflow diameter and system feed rate are the crucial parameters and dictate the
flotation outputs. However, the advantage of the 15-inch cyclone makes the flotation process
more compact thus increasing its potential for capacity per unit volume. Therefore, these
Table 4.5. Evaluation of Pressure Requirements for 15” Flotation Cyclone.
Inputs Results
ΔP
D (in) Q (GPM) D (in) Q (scfm) Q (GPM) Q (GPM) H (in)
uf s p g u o (psi)
1 40 2 2 9.07 8.87 6.5 1.58
1.25 40 2 2 10.72 7.65 6 1.71
1.5 40 2 2 13.55 5.84 5.5 1.64
2 40 2 2 20.58 0.85 4.5 1.89
2.5 40 2 2 22.68 0.00 3.5 0.39
1.5 30 2 2 5.79 0.00 1 0.58
1.5 40 2 2 6.53 0.47 4 1.89
1.5 50 2 2 8.12 2.48 6.5 1.89
1.5 60 2 2 8.38 3.83 8.5 1.89
58 |
Virginia Tech | conclusions led to the use of the 15-inch cylinder length in the coal flotation analysis of the
secondary design.
In addition, the operational parameters of the secondary flotation cyclone were evaluated
using water as the testing media. Table 4.6 represents the testing of the operational parameters
feed rate (gpm), frother concentration (ppm), and gas flow rate (scfm) and their associated
effects on tangential underflow, axial underflow, and overflow. The purpose of testing the
operational parameters was to identify the conditions that controlled the overflow product and
achieve a system that mimics a traditional operating flotation process. The goal was to develop a
central froth column and limit the amount of water reporting to the overflow, which in flotation
Table 4.6. Operating Parameter Testing of Secondary Flotation Cyclone.
Operating Variables Measured Results
Slurry Gas Tangent U/F Axial U/F Overflow Q
s
Flow Flow Frother
Rate Rate (ppm) GPM % Feed GPM % Feed GPM % Feed GPM
(GPM) (scfm)
40 1 5 17.52 0.60 9.39 0.32 2.40 0.08 29.31
40 1 10 18.51 0.63 8.34 0.28 2.70 0.09 29.56
40 1 15 15.45 0.54 8.48 0.29 4.91 0.17 28.84
40 1.5 5 19.86 0.68 8.30 0.28 1.14 0.04 29.30
40 1.5 10 17.70 0.62 8.06 0.28 2.57 0.09 28.33
40 1.5 15 15.46 0.56 8.13 0.29 4.16 0.15 27.74
50 1 5 22.95 0.66 9.64 0.28 2.33 0.07 34.92
50 1 10 24.65 0.67 8.93 0.24 3.18 0.09 36.76
50 1 15 21.47 0.61 8.37 0.24 5.58 0.16 35.42
50 1.5 5 23.16 0.68 8.52 0.25 2.48 0.07 34.16
50 1.5 10 21.86 0.64 8.99 0.26 3.32 0.10 34.17
50 1.5 15 20.04 0.57 9.11 0.26 5.95 0.17 35.11
60 1 5 30.60 0.73 9.05 0.21 2.44 0.06 42.10
60 1 10 27.68 0.65 9.96 0.23 5.12 0.12 42.76
60 1 15 22.69 0.54 9.86 0.23 9.84 0.23 42.39
60 1.5 5 30.86 0.74 8.01 0.19 2.56 0.06 41.42
60 1.5 10 25.10 0.61 11.07 0.27 5.10 0.12 41.27
60 1.5 15 23.22 0.56 11.04 0.27 7.05 0.17 41.31
59 |
Virginia Tech | can increase the product ash content through entrainment. Achieving these conditions allow
optimal conditions for flotation in secondary flotation cyclone design.
Analysis of the operating parameters will indicate the critical parameters that can be
changed to achieve the desired product in addition to verifying the system acts like a flotation
process. The correlation coefficients signify frother concentration has the largest positive impact
on the product overflow (Table 4.7). The feed input flow rate positively impacts all the product
streams with the tangential underflow subjected to the largest impact. The increase in gas flow
rate has a small negative impact on all three product streams, but significant testing could not be
achieved as excessive air quantities “burped” the flotation cyclone. In addition, it should be
noted the presence of the Cavitation Tube reduces the inlet feed flow rate due to the decrease in
pressure and the generation of bubbles. This phenomenon does not affect the slurry rate
measured by the flow meter, but instead may reduce the centrifugal force developed within the
cyclone and inhibit the development of the air core.
4.2.2 Coal Flotation
The evaluation of the secondary design using coal as the flotation feed dictated whether
Table 4.7. Correlation of Operating Inputs and System Outputs.
Input Output Correlation
T U/F (GPM) 0.858
Q s (GPM) A U/F (GPM) 0.616
O/F (GPM) 0.464
T U/F (GPM) -0.053
Q g (scfm) A U/F (GPM) -0.048
O/F (GPM) -0.107
T U/F (GPM) -0.412
C f (ppm) A U/F (GPM) 0.163
O/F (GPM) 0.791
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Virginia Tech | the design could be an effective flotation technique. After optimizing the design and operating
parameters of the flotation cyclone, a coal flotation analysis on the sample was performed and
analyzed. Table 4.8 shows the input operating and design parameters with the feed, product, and
tailing stream characteristics.
The initial results of the secondary design’s capability of achieving a quality flotation
product provide significant information. The design achieved a product with an ash of 20.5 %
from a 47.33% ash feed. However, the design did not significantly affect the ash content of the
tailings stream as the content was only increased to 47.71%. However, this test was performed
incorrectly as the air flow rate was not injected with any pressure. The air flow rate was
established before the test began and equilibrium was created within the pipe network where
Table 4.8. Initial Flotation Analysis of Secondary Design.
Cyclone Design Parameters
Flotation Height 5.25 in
Pedestal Diameter 2 in
O/F Diameter 1 in
U/F Diameter 1 in
Tangential U/F 25 GPM
Operating Parameters
Feed Rate 41 GPM
Air Flow 7 SCFM
Frother Concentration 7 ppm
Collector Concentration 0.065 lb/ton
Stream Characteristics
Product 20.46 % Ash
Axial U/F 51.47 % Ash
Tangential U/F 47.71 % Ash
Feed 47.33 % Ash
Flotation Analysis
Mass Yield 7.35 %
Product Ash Content 20.46 %
Combustible Rec. 11.09 %
61 |
Virginia Tech | minimal air, not 7 SCFM, was injected. Therefore, another test was performed.
The following flotation test was performed with the proper air addition corrections. Table
4.9 shows the design and operating parameters, stream characteristics, and flotation analysis of
the experiment. The analysis did not show any improved qualities in the product after correcting
the air flow rate and increasing the frother and collection addition. These results are particularly
due to a majority of the feed reporting to the underflow streams in a short time period. Therefore,
large fractions of the initial coal weight are not recovered in that short time period and leads to a
tailings composition similar to the feed. Therefore, it was concluded to reduce the volumetric
flow rates of the underflow streams, specifically the tangential underflow, in order to either
decrease the required feed rate or increase the overflow flow rate.
Table 4.9. Secondary Flotation Cyclone Performance with Corrected Air Injection.
Cyclone Design Parameters
Flotation Height 12.5 in
Pedestal Diameter 2 in
O/F Diameter 1.25 in
U/F Diameter 1 in
Tangential U/F 26.4 GPM
Operating Parameters
Feed Rate 50 GPM
Air Flow 1 SCFM
Air Pressure 25 psi
Frother Concentration 10 ppm
Collector Concentration 0.2 lb/ton
Stream Characteristics
Product 23.81 % Ash
Axial U/F 52.74 % Ash
Tangential U/F 47.52 % Ash
Feed 47.15 % Ash
Flotation Analysis
Yield 5.85 %
Ash 23.81 %
Combustible Recovery 8.43 %
62 |
Virginia Tech | In efforts to reduce underflow volumetric flow rates and increase the overflow, new
optimization procedures were taken. The tangential underflow was reduced to zero and the
frother concentration was increased to 30 ppm. The design and operating parameters were
optimized to produce a desirable flotation product using water and the parameters were
evaluated. The results from this evaluation are summarized in Table 4.10.
The results from the third test are promising as the product quality was greatly improved
over the previous tests by rendering the tangential underflow non-existent and increasing the
frother and collector dosages. In order to indicate performance, this point was plotted on the ash
recovery curve of the coal release analysis and compared with the flotation column at the plant
(Figure 4.2). The concentrate ash content was reduced to 4.41% which signifies a high quality
Table 4.10. Flotation Results with Axial Underflow.
Cyclone Design Parameters
Flotation Height 11.25 in
Pedestal Diameter 3.5 in
O/F Diameter 1.25 in
U/F Diameter 2.5 in
Operating Parameters
Feed Rate 50 GPM
Air Flow 0.5 SCFM
Air Pressure 60 psi
Frother Concentration 30 ppm
Collector Concentration 0.25 lb/ton
Stream Characteristics
Product 4.41 % Ash
Axial U/F 61.29 % Ash
Feed 47.16 % Ash
Flotation Analysis
Mass Yield 24.84 %
Product Ash Content 4.41 %
Combustible Rec. 44.94 %
63 |
Virginia Tech | Figure 4.2. Comparison of Flotation Cyclone and Release Analysis.
product, and the design achieved a yield of nearly 25% and recovered nearly 45% of the coal. By
implementing a one tailing discharge system, the axial tailings were downgraded to a 61% ash
content which show significant improvements over the previous system where the feed and
tailings ash were essentially the same. However, a frother concentration of 30 ppm was need to
achieve the product which follows Miller’s ASH testing but greatly exceeds the frother dosage of
the plant column. Although these results show the systems capability to produce a high quality
concentrate, the yield and recovery rates for a single cycle are not comparable to traditional
flotation techniques and typical frother concentrations.
In efforts to increase the yield and combustible recovery of the secondary flotation
cyclone design, a continuous flotation process was utilized. The concentrate was collected at two
minute intervals and the tailings returned to the flotation cyclone for further separation. Prior to
sample number six and eight, additional frother was added to return the frother concentration to
30 ppm and increase recovery of coal. The results of this test are in Table 4.11 and the
comparison to the release analysis in Figure 4.3. The analysis shows the requirement to collect
64 |
Virginia Tech | acts in a similar behavior of flotation columns and conventional cells as large retention times are
required. However, the unit residence time for the second flotation design is 0.66 seconds.
4.3 Tertiary Design: Tangential Aeration Chamber with Axial U/F
4.3.1 Water Only
Analyzing the third flotation cyclone design through the use of water and frother
provided the conclusion that the design was incapable of producing a froth product. Major
problems arose from the attempts to optimize the design and operating parameters to mimic
flotation behaviors. The aeration cylinder’s inability to produce bubbles caused the added air of
the system to short circuit to the overflow and water filled up the aeration chamber. In order to
solve the problem of water filling the aeration chamber, changes were made to increase the
volume of the air without drastically increasing the pressure. However, increasing the volume
resulted in the increase of air short circuiting to the overflow. To overcome the shortcoming of
the bubble generation problem, frother concentration was added to further reduce the surface
tension of the water and the feed rate was increased in efforts to increase the shear rate of the
swirling water against the incoming air. Increasing the frother had no effects and increasing the
feed rate resulted in the aeration chamber filling up faster. These results show this designs
inability to produce the desired froth column mainly due to the aeration cylinder not being an
adequate sparging system.
Due to the conclusions of the initial design, the variation to include a static mixer as an
external sparging system was made. Through optimization of the design and operating
parameters, the third flotation cyclone design was able to achieve a desired flotation behavior.
Upon observation with water, bubbles entered the outer aeration chamber, entered the inner
cylinder, and agglomerated in the center of the cyclone. Subsequently when air and frother was
66 |
Virginia Tech | Table 4.12. Design and Operating Parameters of Third Flotation Cyclone Design.
Design Parameters
U/F Diameter 2 in
O/F Diameter 1.25 in
Pedestal Diameter 1.5 in
Flotation Height 13 in
Operating Parameters
Feed Rate 19 GPM
Air Flow Rate 0.5 SCFM
Frother Concentration 60 ppm
Air Pressure 20 psi
U/F Pump Speed 87 %
added, the concentration of bubbles increased and the air core diameter increased. These
parameters are listed in Table 4.12. These parameters achieved a froth column in which an
adequate overflow product with minimal water was generated. The high frother concentration
increased the diameter of the froth column inside the two inch cylinder and the low air flow rate
provided the proper air to water mixture passing through the static mixer to prevent burping and
large bubbles.
4.3.2 Coal Flotation
Using the design and operating parameters determined in the water testing of the third
flotation cyclone design, a continuous flotation process using the metallurgical coal sample was
performed. A feed and concentrate sample was collected for five second periods once the process
reached steady state and the tailings were reprocessed. Concentrate samples were collected at
five minute intervals and the results are shown in Table 4.13.
The third design resulted in a higher recovery and yield than the previous two designs
while producing an acceptable ash content value. The composite ash value is 11.32 % while the
rejected tails is approximately 37.8% ash. Using the typical flotation performance factors, the
67 |
Virginia Tech | Table 4.13. Continuous Process of Third Flotation Cyclone.
Individual (%) Cumulative (%)
Mass Ash Mass Ash Yield Recovery
C1 10.57 9.86 10.57 9.86 11.6 14.1
C2 6.2 14.13 16.77 11.44 18.4 21.9
C3 9.03 10.65 25.8 11.16 28.4 33.8
C4 7.9 11.59 33.7 11.26 37.1 44.2
C5 8.25 11.54 41.95 11.32 46.1 54.9
Tails 48.99 37.79 90.94 25.58 100.0 99.9
Feed 128.83 25.54
cumulative yield and recovery of the flotation cyclone after a 20 minute operation are 46.1% and
54.9% respectively. At a feed rate of 19 gallons per minute and a 60 ppm frother concentration,
the flotation cyclone design acts more like a traditional flotation cell rather than a fast flotation
process. The required retention time and high frother concentration to achieve these results
defeats the objective of this research project. The third flotation design requires such high
retention times because the mass flow rate of hydrophobic particles through the overflow is such
a small rate on the individual level. In addition, the high frother concentration is necessary to
develop a substantial froth column in the center of the cyclone and provide sufficient bubbles for
particle attachment.
An adequate comparison between the third flotation cyclone design and the industrial
plant data is difficult to conclude as the feed for the third design contained far less ash by weight
than the original industrial coal sample. The original samples from the plant were approximately
47% ash while the flotation cyclone feed ash content was approximately 25%. The cause of this
discrepancy is uncertain. One possible solution is the through flotation cyclone testing, the
sample could not be transferred between the sump and sample containers. This inability to
recover the complete sample from the sump could credit the loss of the large majority of the ash
68 |
Virginia Tech | 5.0 CONCLUSIONS AND RECOMMENDATIONS
In the evaluation of the three flotation cyclone designs, the ability to develop a flotation
device was a challenging task. The initial flotation cyclone design had an inherent disadvantage
of the stationary pedestal. This design flaw created the possibility of bubble particle
agglomerates missing the collection zone and reporting to the underflow. Despite this design
flaw, water testing validated the theory of centrifugal flotation through the creation of an air core
with an external sparging system.
The second flotation design incorporated the design on an adjustable pedestal and
improved the design by adding an axial underflow to control the fluid level within the cyclone.
Evaluation of the design with water eliminated the need for a tangential underflow but proved
the axial underflow did not add a source of fluid control. The coal flotation proved the design
achieved a very high quality product but at a low recovery rate. By evaluating the design as a
continuous process, the system was able to achieve combustible recoveries nearing 73% at ash
contents of 13.7%. However, the design required 16 minutes of reprocessing tailings to achieve
these values thus making the design behave more like a typical flotation technology as opposed
to a fast flotation technology. In addition to the long retention time, increasing the aeration rate
was not a feasible option to increase flotation which adds a limitation to the design.
The third flotation design strayed from the original two designs by incorporating a
tangential aeration chamber and an external static mixer as the bubble generator. The system was
able to achieve an 11.32% ash concentrate at a recovery of 55% in a 20 minute retention time.
Further analysis of the third design as a continuous process didn’t yield higher recoveries.
Although the recovery rate and ash values portray effectiveness, the majority of the feed was
floatable material and this may have led to the increased recovery rate. In addition, the third
70 |
Virginia Tech | design required higher frother concentration not typically found in flotation systems and
potentially could have contributed to the better performance.
The development of a centrifugal flotation technology proved separation results but not to
the standard of traditional flotation techniques. The major limiting factors of the three designs
was controlling the level within the cyclone, providing adequate aeration with an external
sparging system, and increasing mass flow rate of hydrophobic particles to the overflow in a
short time frame. In regards to these three designs, further work could possibly be performed
with flotation feed containing a low concentration of hydrophobic particles with the second
flotation design. The design proved the ability to effectively recover high quality particles at low
yield rates, but no tests were performed to study the size of the recovered particles. In addition,
the Cavitation Tube was designed for flow rates that far exceeded the needs of the second design.
Possible improvements to recovery could be made by utilizing a Cavitation Tube able to handle
those flow rates. In regards to improving the designs and the movement towards fast flotation
technologies, the initiatives should be towards developing an appropriate sparging method that
can be implemented, withstand plugging, and achieve high recovery rates.
71 |
Virginia Tech | EVALUATION OF NOVEL FINE COAL DEWATERING AIDS
by
Mert K. Eraydin
R.H. Yoon, Chairman
Mining and Minerals Engineering
(ABSTRACT)
The costs of cleaning fine coal are substantially higher than those of cleaning
coarse coal. Therefore, many many coal companies in the U.S. choose to discard fine coal
(150 micron x 0) by means of 6-inch diameter hydrocyclones. The cyclone overflows are
stored in fine coal impoundments, which create environmental concerns and represent
loss of valuable national resources. The major component of the high costs of cleaning
fine coal is associated with the difficulty in fine coal dewatering. Therefore, the
availability of efficient of fine coal dewatering methods will greatly benefit companies
In the present study, three different novel dewatering aids have been tested. These
include Reagents W (RW), Reagent U (RU), and Reagent V (RV). These reagents are
designed to increase the contact angles of the coal samples to be dewatered, which should
help decrease the Laplace pressure of the water trapped in filter cake and, hence, increase
dewatering rate. They were tested on i) the fresh coal samples from Consolidation Coal
Corporation’s Buchanan Preparation Plant, ii) a composite drill core sample from the
Smith Branch Impoundment, Pinnacle Mine Mining Company, and iii) a blend of coals
from the Smith Branch Impoundment, thickener underflow, and thickener feed.
The coal samples were used initially for laboratory-scale tests using a 2.5-inch
diameter Buchner vacuum filter. The results showed that the use of the novel dewatering
aids can reduce the cake moisture up to 50% over what can be achieved without using
any dewatering aid. The use of the dewatering aids also increased the kinetics of
dewatering by up to 6 times, as measured by cake formation times.
On the basis of the laboratory test results, pilot-scale continuous vacuum filtration
tests were conducted using a 2-feet diameter Peterson vacuum disc filter. The cake |
Virginia Tech | ACKNOWLEDGEMENTS
With utmost appreciation, I would like to thank my advisor Dr. Roe-Hoan Yoon,
for his guidance, enthusiasm, support and patience exhibited throughout the course of my
study. Special thanks are also extended to Dr. Jerald H. Luttrell, and Dr. Ismail Yildirim,
for their guidance and specific help with experiment and data analysis. I would also like
to than the committee members, Dr. Greg Adel and Dr. Tom Novak, for their
encouragement and useful suggestions. I am also grateful to Dr. Jinming Zhang, Mr.
Todd Burchett, Mr. Christopher Barbee, Mr. Selahattin Baris Yazgan, and Mr. Emilio
Lobato for their friendship and support.
I owe special gratitude to Mr. Hubert C.R. Schimann for his friendship, helpful
suggestions and continuous support.
I would like to express my most sincere appreciation to my parents Ziynet and Ali
Eraydin for their inspiration, encouragements, moral and financial support and continued
suggestions. I would also like to thank my sister Feyza for her continued support.
Finally I would like to express my thanks to Ayse Ceren Talu for all her
encouragements, patience and dedication.
iv |
Virginia Tech | INTRODUCTION
Coal preparation plants consist of simple crushing, sizing and cleaning operations
for coarse and intermediate coal size fractions but commonly involve complex cleaning
circuits for fine size fractions which also require dewatering processes. These can be
broadly classified into three groups; sedimentation, filtration and thermal drying. Since
the market demand is increasing for low-sulfur, high grade coal as a result of economic
and environmental considerations it causes processes to be extended towards the ultra
fine range which accompanies an increase in the yield of ultra fine tailings. Ultra fine
tailings are particularly difficult to dewater because of their large surface area. The
performance of most types of dewatering equipment depends strongly on several
parameters of particle-aqueous systems. Knowledge of these parameters would help the
processor to make suitable modifications for achieving improved filtration and
dewatering. These parameters include mineral type, particle size and distribution, surface
characteristics of the materials, applied pressure, cake thickness, clay content, surface
oxidation, solid /liquid ratio and presence of chemical additives.[1-6]
The coal preparation plants, in which the fine coal fraction is processed, utilize
filtration technologies that typically include vacuum disc/drum filters, plate and frame
filter presses and horizontal belt filter for dewatering. Filtering is a solid-liquid separation
process where solid particles are separated from liquid by forcing the slurry to pass
through a suitable filtering medium with the help of either vacuum or pressure force. This
allows only the liquid to pass through, leaving the solid particles to form a “cake”. As the
final step or when the efficiency of dewatering becomes so low, dewatered product is
thermally dried in thermal dryer units to obtain the required moisture contents. [1, 2, 4-9]
To avoid the expensive and environmentally important thermal drying, efforts are
being made in the coal preparation sector to counteract this trend both by improvement
and further development of already known dewatering methods and equipment and
1 |
Virginia Tech | testing those not previously used in coal preparation. The availability of efficient
processes that can reduce the moisture of fine particles will greatly benefit companies if
these processes can be applied to reduce the moisture content of fine coal. [1, 6, 9-13]
To better understand the problems associated with the the process of fine particle
dewatering and to study the effects of the parameters affecting the process, bench-scale
tests were conducted on fine coal samples. The novel dewatering aids, namely RU and
RW were tested. Pilot-scale filtration tests were conducted on fine coal samples, and the
results were compared with the the laboratory test results. It was found that the use of the
novel dewatering aids can substantially reduce the cake moistures by up to 50%. The role
of the reagents is to increase the hydrophobicity (or dewettability) of the coal. In the
presence of these reagents, the contact angles of the coal samples increased, preventing
the surface of the particles to be wetted by water molecules. It is suggested that the
increase in contact angle is responsible for the reduction of the capillary pressure in a
filter cake, which in turn should reduce the cake moisture. The novel dewatering aids
also cause a decrease in the surface tension of water and increased cake porosity (or
capillary diameter), both of which should contribute to lowering cake moistures. [1, 6, 9-
16]
REFERENCES
1. Asmatulu, R., Advanced Chemical-Mechanical Dewatering of Fine Particles, in
Material Science and Engineering. March 12, 2001, Virginia Polytechnic Institute
and State University: Blacksburg.
2. Wills, B.A., Coal preparation technology: D.G. Osborne Graham & Trotman
Ltd, Sterling House, 66 Wilton Rd., London SW1V 1DE. 1988257.50. 1175 pps. (2
vols) ISBN 185333 092 2. Minerals Engineering, 1988. 1(4): p. 374-375.
3. Ottley, D.J., Mineral processing technology (4th edition): B.A. Wills Pergamon
Press, Oxford, 1988. 785 pps. ISBN 0-08-034937-4 (hardcover); ISBN 0-08-
034936-6 (flexicover). Minerals Engineering, 1988. 1(3): p. 268-269.
4. Osborne, D.G., Vacuum Filtration-Part 1. p. 415-433.
5. Osborne, D.G., Rotary-vacuum filtration of coal flotation concentrates.
International Journal of Mineral Processing, 1976. 3(2): p. 175-191.
2 |
Virginia Tech | ABSTRACT
Three different dewatering reagents were tested, coded as RW, RU and RV, on
Consolidation Coal Corporation’s Buchanan Preparation Plant’s for the fine coal
dewatering circuit (28 mesh x 0). The role of these reagents was to increase the
hydrophobicity of the coal for dewatering. According to the Laplace equation, an increase
in contact angle with the surfactant addition should decrease the capillary pressure in a
filter cake, which should in turn increase the rate of dewatering and help to reduce the
cake moisture. The use of the novel dewatering aids causes a decrease in the surface
tension of water and an increase in the contact angle and porosity of the cake, both of
which also contribute to improved dewatering. It was found that final cake moistures
could be reduced by 20-40% by adding novel dewatering chemicals. It was also
determined that the use of the novel dewatering aids could reduce the cake formation
time by a significant degree due to the increased kinetics of dewatering. The Buchanan
coal (Pocahontas seam) contains a large amount of calcium chloride. The Ca2+ ions
present in the froth product cause coagulation of fine particles, which in turn traps
moisture within the flocs that are formed. Several series of laboratory dewatering tests
were conducted to determine if water chemistry was indeed a problem in dewatering and,
if so, whether a viable solution could be found to eliminate or minimize the problem. The
effects of various water treatment strategies were evaluated. These treatments included
the addition of acid/base (to control pulp pH), sodium carbonate (Na CO ),
2 3
ethylenediamine tetraacetic acid (EDTA), sodium silicate (Na SiO ) to precipitate Ca2+
2 3
ions, oxalic acid, succinic acid, ammonium oxalate, Na-hexametaphoshate, calcium oxide
and hydrogen peroxide (H O ). The results showed that the use of sequestrating reagents
2 2
for water treatment in conjunction with dewatering aids reduced the cake moistures by a
greater percentage than by using the reagents alone, the extent of which depend on the
particle size, cake thickness, drying time, reagent dosage, conditioning time, reagent type,
5 |
Virginia Tech | INTRODUCTION
In general, upgraded fine coal by flotation contains approximately 80% water. To
begin the dewatering process, the slurry can be thickened to 35 to 75% in large settling
tanks. For further dewatering, the product is subjected to a mechanical dewatering
process such as filtration. Filtering is a solid-liquid separation process where solid
particles are separated from liquid by forcing the slurry to pass through a suitable filtering
medium with the help of either vacuum or pressure force. This allows only the liquid to
pass through, leaving the solid particles to form a “cake”. As a final step, the dewatered
product is thermally dried in thermal dryer units to obtain the required moisture
contents[1-5].
In the absence of a fine coal cleaning circuit, a majority of coal preparation plants
use screen bowl-centrifuge to dewater 1mm x 150 micron size clean coal. But the cost of
the dewatering fines sharply increases when the particle is below 150 microns. For this
reason it is often more economical to discard the minus 150 micron size fraction of the
plant’s run of mine coal without any attempt to recover the clean coal content of this size
fraction if it is only a small fraction (5 to 10%) of the product stream. The coal
preparation plants, in which the fine coal fraction is processed, utilize filtration
technologies that typically include vacuum disc/drum filters, plate and frame filter
presses and horizontal belt filter for dewatering.[1, 2, 5-8]
The performance of most of the dewatering equipment that is used today depends
strongly on several parameters of particle-aqueous systems. Knowledge of these
parameters would help the processor to make suitable modifications for achieving
improved filtration and dewatering. These parameters include mineral type, particle size
and distribution, surface characteristics of the materials, applied pressure, cake thickness,
clay content, surface oxidation, solid /liquid ratio and presence of chemical additives.[1,
9-11]
7 |
Virginia Tech | FILTRATION THEORY
Cake filtration and dewatering has been characterized over the years via the
capillary pressure function, which is the relationship between the pressure and saturation.
In this characterization, it is assumed that the filter cake consists of a number of
capillaries of varying diameters. Such a filter cake is saturated with liquid initially and
requires a minimum threshold pressure drop for the start of drainage of liquid from cake
itself. One can use the Laplace – Young Equation, given below, as the most simple and
predictive guide to characterizing the drainage of liquid from the capillary of radius `r`.
[1, 15]
2γ cosθ
[1]
∆P = LV LS
r
where γ is the liquid surface tension and θ is the solid-liquid contact angle. In the case
LV LS
of a constant pore size distribution, the threshold pressure (∆P ) can approximately be
b
estimated by r = r . For estimation of threshold pressure drop for a bed of packed
max
spheres, Carman (1946) suggested an expression which was later used to define modified
threshold pressure drop (∆P *). This in turn was later used to characterize dewatering
b
properties of alumina trihydrate.
k(1−ε)γ cosθ
∆P* ≅ ∆P= LV LS [2]
b εd
p
where ε is the cake porosity, d is the specific mean surface diameter and k is a constant,
p
the value for which was 6.0 (for the case of packed spheres) as estimated by Carman
9 |
Virginia Tech | (1946). Experiments conducted by Rushton and Wakeman (1977), showed that the value
of `k` lies between 4.6 and 8.0.[1, 15, 16]
The process of draining liquid from a filter cake can be seen as consisting of three
stages. At the beginning of a dewatering cycle, or the first stage, the drainage of liquid
from the cake does not occur as long as ∆Pis lower than ∆P , and the cake is said to be
b
completely saturated (S=1.0). This means that all the pore spaces of the cake are
completely filled with liquid. At pressures≥∆P , saturation falls below 1.0 and
b
eventually approaches irreducible saturation (S ) at pressure∆P . A further increase in
∞ ∞
vacuum will not lower the limit of saturation. The pressure∆P could be viewed as the
∞
capillary pressure of minimum drainable pore (r ). For I intermediate pressures, there is
min
a steady decrease in moisture content, and the cake will achieve an equilibrium saturation
(S ).It was defined the residual saturation (S ) of cake by the following expression.
p R
S −S
S = p ∞ [3]
R 1−S
∞
A number of empirical relations have been proposed by various researchers
(Wakeman, 1975, 1976; Wainwright, 1986) to correlate the applied pressure differential
and cake saturation.
ξ
∆P*
S = b [4]
R
∆P
where ξis the pore size distribution index and its value lies between 2 and 8. The
residual saturation is related to the moisture content of the cake by the following
relation.[1, 15]
ε ρ
S l 100%
Moisture Content = [5]
R 1−ερ
s
10 |
Virginia Tech | EXPERIMENTAL
The objectives of this study were to i) identify the best possible reagents and
combinations thereof for this specific coal, ii) identify the conditions under which a given
dewatering aid can give the best performance, and iii) solve water chemistry problems
caused by the Ca2+ ion. To meet these objectives, a series of laboratory dewatering tests
were conducted by using novel dewatering aids, namely RW, RV, and RU, which were
developed at Virginia Tech, and by using various sequestrating reagents such as EDTA,
Na SiO , oxalic acid, succinic acid, ammonium oxalate, sodium-hexametaphoshate,
2 3
calcium oxide, and hydrogen peroxide to control the water chemistry on Buchanan coal
samples. Furthermore, a detailed study was performed to understand the effects of
parameters such as pH, filter medium, rinsing/washing, desliming, cake thickness,
mixing, cleaning and re-cleaning of samples under laboratory conditions. A few tests
were also conducted using samples of aged coal to study the effect of oxidation on the
performance of dewatering reagents.
The laboratory test work was conducted on three different samples taken from
Consolidation Coal Corporation’s Buchanan Preparation Plant in Mavisdale, Virginia.
These samples were i) a flotation feed sample, ii) a grab sample of current flotation
product, and iii) a slip-stream sample of filter feed (all taken on various dates). Unless
otherwise stated, the dewatering tests were conducted within three days of receipt of
sample in order to minimize coal oxidation effects.
When the plant flotation feed sample was used in the dewatering tests, the sample
was first floated in a laboratory mechanical flotation cell using 300 g/t of kerosene and
150 g/t of MIBC to remove ash-forming minerals. The floated product was then subjected
to the dewatering tests at about 20% solids. Reagents RW and RU, diluted to 33.3% in
solvent, were used in these tests. Immediately prior to each filter test, a known volume of
slurry was conditioned with the reagent in a mechanical shaker and then poured into a
11 |
Virginia Tech | Buchner funnel before applying vacuum. The following conditions were kept constant
during the tests:
(cid:131) Vacuum:….……………68 kPa (20 inches Hg)
(cid:131) Drying cycle time………….………2 minutes
(cid:131) Cake thickness…………………….10-15 mm
(cid:131) Volume of slurry………………………100ml
(cid:131) Reagent conditioning time..………..5 minutes
A limited number of standard analyses (e.g., particle size analysis, ash analysis)
were performed on selected samples to provide baseline data related to this particular
sample.
For laboratory-scale batch dewatering tests, the samples were collected either in
5-gallon buckets or 55gallon drums. To be able to receive a representative sample, slurry
samples were homogenized by mixer. While they were being agitated, a volume of the
slurry was scooped out of the 5-gallon bucket in a dipper of known volume and used for
the dewatering studies.
A Buchner funnel, a ¾ HP Vacuum Pump and filter cloths of different porosities
and diameters were used as laboratory-scale filter equipment. The Buchner funnel was
used to test different dewatering and sequestrating chemicals in these experiments. For
this particular study, all tests were conducted using a 2.5-inch diameter Buchner funnel
with filter cloth and wire screen mesh supplied from Peterson Co. The height of the
funnel was 8-inches. In a given experiment, a known volume of coal slurry was poured
into the funnel and vacuum was applied to form a cake on the filter medium. Since the
diameter of the funnel is fixed at 2.5 inches, it determines the cake thickness.
Figure 1 shows the apparatus used for the Buchner funnel filtration tests. A
Buchner funnel was mounted on a vacuum flask, which in turn was connected to a larger
vacuum flask to stabilize the vacuum pressure. A known volume of a slurry sample was
poured into the funnel before opening the valve between the two flasks in order to subject
the slurry to a vacuum pressure. In each experiment, the cake thickness, applied and
actual vacuum pressures and cake formation time were recorded.
12 |
Virginia Tech | VACUUM
TO VACUUM
PUMP
BUCHNER
FUNNEL
VALVE
VACUUM FLASK VACUUM FLASK
Figure 1 - Experimental setup for laboratory vacuum filtration tests
A known volume of coal or mineral slurry was transferred to a 250-ml flask, to
which a known amount of a dewatering aid (or a mixtures thereof) was added by means
of a Microliter syringe. The flask was attached to a mechanical shaker or high intensity
motor mixer for a given period of time to allow for the reagent to adsorb on the surface of
the particles to be dewatered. After the conditioning time, the slurry was transferred to
the Buchner funnel. Filtration was commenced when a vacuum was applied to the slurry.
The time it took for the bulk of the water to pass through the filtering medium is referred
to as cake formation time. After cake formation, the vacuum pressure was maintained for
a desired period of time to remove the residual water trapped in the capillaries formed
between the particles in the cake. After allowing the reagent to be adsorbed on the surface
of particles, sequestrating chemicals were added and agitated for a given time. After this
period, which is referred to as drying cycle time, part of a representative filter cake was
removed, weighed, and dried overnight in a conventional oven. The sample was weighed
again after drying, and the moisture content was calculated from the difference between
the dry and wet weights.
13 |
Virginia Tech | RESULTS AND DISCUSSIONS
1. Effect of Chemicals
Effects of RU and RW
Table 1 gives the laboratory test results obtained on the Buchanan plant flotation
product using RU and RW as the dewatering aids at 68 kPa (20 inches Hg) vacuum. As
shown, the moisture content in the filter cake was reduced with increasing reagent
addition. At RU additions of 0.5 and 2.5 kg/mt, the moisture contents of the cake were
reduced from 17.6% to 16.1% and 14.4%, respectively. Similar results were obtained
with RW as a dewatering aid. These values corresponded to a 15-20% moisture reduction
in the filter product.
Table 1 - Effect of reagent addition on dewatering Buchanan’s flotation product
Reagent Dosage Moisture Content (%wt)
(kg/mt) Reagent W Reagent U
0 17.6 17.6
0.5 16.2 16.1
1.5 16.2 14.9
2.5 15.0 14.4
Similar dewatering tests were conducted on the filter feed sample (which contains
approximately 10 g/t of Nalco 9806 polymer flocculant as the dewatering aid) using RW
and RU as dewatering aids. Results for the filter feed sample are summarized in Table 2.
The results show that the moisture content of the filter product again decreases with
increasing RW and RU additions from 0.5 to 2.5 kg/mt (1 to 5 lb/ton). In this case, the
addition of 2.5 kg/mt (5 lb/ton) of reagent RW reduced the cake moisture from 18.1 to
14.9%, giving a percentage moisture reduction of about 20%. The moisture reduction is
14 |
Virginia Tech | quite similar to that obtained for the flotation product except that the moisture content of
the filter cake product obtained using reagent RU was almost 2 percentage units lower,
i.e., 14.4% vs. 16.6% moisture in the filter cake (see Tables 1 and 2). The reasons for the
relatively poorer behavior of reagent RU may be related to the presence of flocculant in
this particular sample. Apparently, the polymer flocculant has an adverse effect on the
performance of RU during dewatering. Those poor results are also due to the Ca2+ ions
present in Buchanan plant water.
Table 2 - Effect of reagent addition on dewatering of Buchanan’s filter feed
Reagent Dosage Moisture Content (%wt)
(kg/mt) Reagent W Reagent U
0 18.1 18.1
0.5 17.58 17.5
1.5 15.2 16.78
2.5 14.9 16.58
Effect of Soda Ash (Na CO )
2 3
One way of eliminating Ca2+ ions present in Buchanan plant water is to introduce
CO -2 ions. This can be achieved by simply adding soda ash (sodium carbonate) to the
3
feed pulp. To test this approach, dewatering tests were conducted using 0.5 kg/mt of soda
ash in conjunction with Reagents RW and RU. The results are presented in Table 3. In
general, sodium carbonate was not particularly effective when Reagent RW was used as
the dewatering aid. In fact, the moisture values actually increased slightly when soda ash
was added. Conversely, the addition of soda ash improved the performance of Reagent
RU at dosages of 1.5 kg/mt or higher. The filter cake moisture was reduced to a very low
value of just 12.3% when 2.5 kg/t of RU was used in conjunction with soda ash. The use
of soda ash was not effective when lower dosages (0.5 kg/mt) of Reagent RU were used.
The amount of Na CO required to increase moisture reduction was high which probably
2 3
indicates that the amount of Ca2+ ions present in the plant water was high
15 |
Virginia Tech | Table 3 - Effect of Na CO addition at 0.5 kg/mt on dewatering of Buchanan’s flotation product
2 3
Moisture Content (%wt)
Reagent Dosage
RW
RU
(kg/mt)
Without With Without With
Na CO Na CO Na CO Na CO
2 3 2 3 2 3 2 3
0
18.1 17.9 18.1 17.9
0.5
17.6 17.7 17.5 17.5
1.5
15.2 15.7 16.8 15.7
2.5
14.9 15.3 16.6 12.3
Effect of Ethylenediamine Tetraacetic Acid (EDTA)
Ethylenediamine Tetraacetic Acid (EDTA) is a well known complexing agent for
metal ions. As such, this chemical can be used to eliminate the adverse effects of metal
cations by forming complexes with them, i.e., Ca2+. The test results, which are given in
Table 4, showed that the filter cake moisture content was lowered by only about one
percentage point after adding 0.2 kg/mt EDTA along with 0.75 kg/mt of Reagent RV.
Therefore, the addition of ETDA was not effective in controlling the water chemistry
problem.
Table 4 - Effect of using EDTA as the blocking agent for Ca2+ ions on dewatering of
Buchanan’s flotation product
EDTA Dosage
Moisture Content (%wt)
(g/mt)
0 17.5
50 17.3
100 17.1
200 16.7
16 |
Virginia Tech | Effect of Hydrogen Peroxide (H O )
2 2
It is possible that coal particles are coagulated due to the hydrophobicity of the
particles. This process is known as hydrophobic coagulation. Water may be trapped
within the coagula, causing high moisture. One way of preventing hydrophobic
coagulation is to make the coal less hydrophobic. Therefore a series of laboratory
dewatering tests were conducted using hydrogen peroxide (H O ) as an oxidizing agent.
2 2
The test results, which are given in Table 5, indicated that very little improvement in
filter cake moisture (only about one percentage point) was obtained when using H O
2 2
with Reagent RV.
Table 5 - Effect of using H O as oxidizing agent on dewatering of Buchanan’s flotation product
2 2
H O Dosage
2 2 Moisture Content (%wt)
(g/mt)
0 17.5
200 17.4
300 16.9
500 16.6
Effect of Sodium Silicate (Na SiO )
2 3
In order to control the water chemistry problem, a series of dewatering tests were
conducted using sodium silicate (Si O -2 ions) as a sequestering reagent for Ca2+ ions.
2 5
The results are given in Table 6. The dosages of sodium silicate varied from 0.5 to 25
kg/t, while the dewatering aid (RW or RU) dosages were kept constant at 2.5 kg/t. The
addition of sodium silicate generally improved the performance of the dewatering aids up
to a dosage level of 3 g/t. For example, the addition of 3 g/t of sodium silicate improved
the performance of Reagents RW and RU by 0.8 percentage points (i.e., 14.3% vs. 13.5%
moisture) and 2.0 percentage points (15.2% vs. 13.2% moisture), respectively. However,
further addition of sodium silicate above 3 kg/t caused an increase in moisture content.
The results indicate that the use of sodium silicate increases moisture reductions beyond
17 |
Virginia Tech | what can be achieved using dewatering reagents alone. It is believed that sodium silicate
not only acts as the sequestering agent for the Ca2+ ions, but also as the dispersing agent
for the colloidal clay particles that may be attached to the surface of coal particles.It can
be said that Na SiO may remove Ca2+ ions, increase ph and helps dispersion of particles.
2 3
Table 6 - Effect of Na SiO addition on dewatering of Buchanan’s flotation product sample
2 3
Moisture Content (%wt)
Reagent Dosage
(kg/mt)
RW + Na SiO RU + Na SiO
2 3 2 3
Baseline (No reagent) 17.3 17.3
2.5 kg/t dewatering aid 14.3 15.2
2.5 g/t dewatering aid +
14.1 14.6
1g/t Na-Silicate
2.5 kg/t dewatering aid
13.5 13.2
+3 g/t Na-Silicate
2.5 kg/t dewatering aid
15.9 14.4
+ 5 g/t Na-Silicate
2.5 kg/t dewatering aid
16.2 14.4
+ 10 g/t Na-Silicate
2.5 kg/t dewatering aid
16.3 14.5
+25 g/t Na-Silicate
2.5 kg/t dewatering aid
16.6 14.6
+ 50 g/t Na-Silicate
Effect of Oxalic Acid
A series of laboratory dewatering tests were conducted using oxalic acid
(COOH) to determine whether this reagent could sequestrate Ca2+ ions and eliminate the
2
adverse effects of Ca2+ ions on filtration. The test results, which are given in Table 7,
show that the moisture content of filter cake could be reduced from 17.3% to 13.3-13.4%
by using 1-50 g/t oxalic acid along with 2.5 kg/t of Reagent RW. Likewise, a filter cake
with as little as 13.3-13.4% moisture could be obtained by using 5-10 g/t oxalic acid
along with 2.5 kg/t RU as dewatering aid. These data show that the use of oxalic acid in
18 |
Virginia Tech | combination with the dewatering aids is a promising approach for overcoming the water
chemistry problem at the Buchanan plant.
The following mechanism may be proposed for the function of oxalic acid in
dewatering of fine mineral particulates. When oxalic acid is dissolved in water, it will
ionize to -OOC—COO-. The di-carboxyl groups will then react with Ca2+ ions to form
calcium oxalate in water. As a result, the performance of dewatering agent (i.e., Reagent
U or Reagent W) is enhanced as the oxalic acid is used as the sequestering agent for the
Ca2.
Table 7 - Effect of Oxalic Acid addition on dewatering of Buchanan’s flotation product
Moisture Content (%wt)
Reagent Dosage
(kg/mt)
RW + Oxalic Acid RU + Oxalic Acid
Baseline (No reagent) 17.3 17.3
2.5 kg/t dewatering aid 14.3 15.2
2.5 kg/t dewatering aid
13.4 16.6
+ 1 g/ Oxalic Acid
2.5 kg/t dewatering aid
13.3 13.4
+ 5g/t Oxalic Acid
2.5 kg/t dewatering aid
13.2 13.3
+ 10g/t Oxalic Acid
2.5 kg/t dewatering aid
13.4 13.5
+ 15 g/t Oxalic Acid
2.5 kg/t dewatering aid
13.5 13.9
+ 25 g/t Oxalic Acid
2.5 kg/t dewatering aid
13.7 14.0
+ 50 g/t Oxalic Acid
Effects of Succinic acid, Ammonium Oxalate and Na-Hexametaphoshate
As shown in Table 8, the addition of succinic acid, (C H O ) generally improved
4 6 4
the performance of the dewatering aids up to a dosage level of 50 g/t of succinic acid,
19 |
Virginia Tech | where the dosages varied from 1 to 50 g/t, while keeping the dewatering aid RW dosage
constant at 2.5 kg/t. For example, the addition of 10 g/t of succinic acid improved the
performance of Reagents RW by 2.75 percentage points (i.e., 15.3% vs. 13.25%
moisture). The same mechanism proposed above for the oxalic acid holds for the succinic
acid as well.
The results show that the moisture content again decreases with increasing
ammonium oxalate additions from 1 to 25 g/t. In this case, the addition of 25 g/t of
ammonium oxalate in conjunction with Reagent RW reduced the cake moisture from
15.3 to 13.4%, giving a percentage moisture reduction of about 14%.
However, the results obtained by using Na-Hexamatephosphate, indicated that
there was a very little improvement (only about less than one percent) in filter cake
moisture.
Table 8 - Effects of Succinic acid, Ammonium Oxalateand Na-Hexamatephosphate on filtration
of Buchanan’s flotation product
Moisture Content (%wt)
Sequestrating
Vacuum
Reagent RW (5 lb/ton) RW (5 lb/ton)
Pressure RW (5 lb/ton)
Dosage + +
(inHg) +
(g/ton) Ammonium Na-Hexameta
Succinic Acid
Oxalate Phosphate
0 15.3 15.3 15.3
1 14.33 15.01 15.22
3 13.28 14.64 14.97
20 5 13.08 14.5 15.27
10 12.5 13.9 15.82
25 12.8 13.4 15.28
5 0 12 .9 X 14.69
Effects of Calcium Oxide (CaO)
Similar dewatering tests were conducted to improve the performance of the
dewatering aids by using calcium oxide. This reagent will increase the pH and may
20 |
Virginia Tech | precipitate Ca2+ ions as Ca(OH) The test results, which are given in Table 9, showed that
2.
the filter cake moisture content was lowered by about 2.7 percentage points after 10g/ton
CaO along with 2.5kg/t of Reagent RW were added. This may be related to increasing pH
or making the coal surface more basic so that the acidic dewatering aids (RU or RW) will
adsorb on coal surface much more than at neutral pH.
Table 9 - Effects of Calcium Oxide (CaO) on filtration of Buchanan’s flotation product
Moisture Content (%wt)
Reagent
Vacuum
Dosage CaO (g/ton)
Pressure
(kg/ton)
(inHg)
RW in Diesel
5(g/ton) 10(g/ton) 15(g/ton) 25(g/ton) 50(g/ton)
0 17.7 17.7 17.7 17.7 17.7
0.5 15.2 15.35 14.92 15.1 16.02
20
1.5 14.09 13.85 13.97 14.89 15.01
2.5 13.4 13.25 13.67 14.31 14.09
Effect of pH
Table 10 shows the effect of pH on the filter cake moistures in the absence and
presence of a dewatering aid (Reagent RW). In these tests, the pH of the slurry was
adjusted using HCl and NaOH. Without a dewatering aid, the lowest moisture values
were achieved around neutral or slightly acidic pH levels (from 6 to 8). The baseline
moisture contents were increased for both the acidic (pH=4) and alkaline (pH=11)
conditions when no reagent was added. Conversely, cake moistures in the 13.6-13.9%
range were obtained for all pH values tested except for pH 8 (which gave a filter cake
moisture of nearly 15%). This relatively high filter cake moisture corresponded to the pH
at which the lowest baseline moisture was obtained, i.e., when no reagent was employed.
These results suggest that a slightly basic (pH 8) solution may adversely affect the
dewatering aids. Additional tests are needed to verify this observation
21 |
Virginia Tech | Table 10 - Effect of slurry pH on dewatering of Buchanan’s flotation product sample
Moisture Content (%wt)
Slurry Moisture
pH Without With Reduction (%)
Dewatering Aid Dewatering Aid
4 18.74 13.86 26.0
6 18.16 13.83 23.8
8 17.00 14.95 12.1
11 20.62 13.60 34.0
2. Effect of Physical Parameters
Effect of filter medium
Tables 11 and 12 show the results obtained with the flotation feed sample. As
shown, the two series of tests conducted using either a steel wire mesh (40 x60) or filter
cloth as a filter medium do not change the performance of the dewatering aids (RU or
RW). However, the moisture content in the base sample (using no dewatering aid)
changes depending on which filter medium was used. As presented in the tables,
moisture content in the base sample was 19% when using a filter cloth, whereas it is only
17% when using steel wire mesh as the filter medium.
Table 11 - Effect of reagent addition on dewatering of Buchanan’s flotation feed using a steel
wire mesh as the filter medium
Reagent Dosage Moisture Content (%wt)
(kg/mt) Reagent W Reagent U
0 17.0 17.0
0.5 14.6 15.0
1.5 13.8 14.8
2.5 13.4 14.6
22 |
Virginia Tech | Table 12 - Effect of reagent addition on dewatering of Buchanan’s flotation feed sample using
a filter cloth as the filter medium
Reagent Dosage Moisture Content (%wt)
(kg/mt) Reagent W Reagent U
0 19.1 19.1
0.5 14.3 14.8
1.5 13.7 14.4
2.5 13.6 14.2
Effect of flotation in laboratory
As shown in Table 13, a filter product with significantly less moisture (13-14%
moisture) could be obtained when laboratory tests were conducted using flotation feed
samples taken from the plant and floated in the laboratory. This very low moisture could
not be obtained using the plant flotation product.
Table 13 - Effect of flotation on dewatering of Buchanan’s flotation feed
Reagent W Dosage Moisture Content (%wt)
(kg/mt) Flotation Product(1) Flotation Product(2)
0 17.0 17.6
0.5 14.6 16.2
1.5 13.8 16.2
2.5 13.4 15.0
Flotation Product (1): Floated in the laboratory
Flotation Product (2): Plant Flotation Product
The difference in the moisture levels between flotation plant product and flotation
feed samples (floated in the laboratory) appears to be due to water chemistry problems
The Ca2+ ions present in the froth product cause coagulation of fine particles, which in
turn traps moisture within the flocs that are formed. The concentrations of polymer
23 |
Virginia Tech | flocculant and flotation reagents in the plant water are also relatively high due to the
circulation of plant filter effluent to the flotation cells. The circulation may be another
reason why the performance of dewatering chemicals is poorer at the plant site. It is also
interesting to note that the problems associated with the water chemistry are eliminated
when the flotation plant product is brought back to the laboratory and aged for two
weeks. The test data collected to date indicate that cake moistures in the 12-13% range
could be achieved at the plant site if the water chemistry problem is resolved.
Effect of Rinsing/Washing with Tap Water
The next series of dewatering tests were conducted after rinsing/washing the
flotation product with fresh tap water to remove any residual flotation reagents (collector
and frother), other plant chemicals (polymer flocculant), and ionic species (Ca2+ etc.) that
might be present in the flotation pulp. In this series of tests, a 2-liter sample of flotation
product slurry from the Buchanan plant was filtered three times after adding 2 liters of
tap water before each filtration step. After the final filtration step, the solids were
adjusted to 25% solids by weight (which is the same as the Buchanan plant flotation
product) by adding fresh tap water. The slurry was then subjected to dewatering tests
using Reagents RU, RW and RV.
Table 14 shows that the moisture content of the baseline (no dewatering aid)
increased to 23% after the plant flotation product was washed with fresh tap water. The
baseline filter cake moisture content was normally about 18% when no treatment was
used. Although the moisture content of the baseline jumped from 18% to 23%, the
moisture content of the cake was reduced to 13-14% after the addition of dewatering aids.
The results show that the lowest achievable moisture does not change appreciably
regardless of the treatment employed. The only parameter that changes is the baseline
moisture obtained when no dewatering aid is used. It should be noted that moisture
reductions as high as 42% were achieved after rinsing/washing the flotation product with
fresh water. These results verify that water chemistry can significantly impact the final
cake moisture regardless of whether dewatering aids are utilized.
24 |
Virginia Tech | Table 14 - Effect of washing Buchanan’s flotation product on dewatering results using RU, RW
and RV dissolved in diesel at 1:2 ratios
Moisture Content (%wt)
Reagent Dosage
(kg/mt) Reagent U (1:2) Reagent W (1:2) Reagent V (1:2)
0 22.9 22.9 22.9
0.5 15.99 15.80 15.70
1.5 15.10 14.08 13.50
2.5 14.45 14.13 13.37
Effect of Aging Slurry
In the next series of tests, laboratory scale dewatering tests were conducted after
aging the Buchanan plant flotation product for three weeks. The dewatering aid (Reagent
RV) was added in pure form (no solvent) and was added as 1:1 and 1:2 ratios of Reagent
RV dissolved in diesel solvent.
Table 15 gives the results of dewatering tests on the aged flotation product from
the plant. As shown, the moisture content obtained in the baseline test (with no
dewatering aid) was about 20%. This value was reduced to almost 13% using 1.5 kg/mt
of dewatering aid and to a very impressive value of almost 12% using 2.5 kg/mt of
dewatering aid. These results show that the performance of the dewatering aid can be
substantially improved by simply aging the coal slurry. These results also indicate that
the performance of Reagent RV is better when dissolved in a solvent (i.e., diesel).
25 |
Virginia Tech | Table 17 - Effects of Aeration on the filtration of Buchanan’s flotation product
Moisture Content (%wt)
Reagent Dosage
Reagent W Reagent W
(kg/mt)
Aged sample As Is
0 18.1 17.6
0.5 15.7 16.2
1.5 14.1 16.2
2.5 13.17 15.0
Effect of Desliming
Several series of dewatering tests were conducted after desliming the filter feed
slurry. The desliming step reduces the amount of ultrafine particles and may also remove
unwanted ions that adversely impact the filter’s performance. Particle size analysis
conducted on the Buchanan plant flotation product showed that it contained a relatively
high proportion (28-30% by weight) of ultrafine minus 400 mesh (minus 38 micron)
material. The desliming was achieved by simply screening the sample at 400 mesh using
a laboratory sieve. The undersize material that passed through the screen was added back
to the oversize product retained on the sieve in the desired proportions. Three series of
tests were conducted in which none (0%), all (100%) and half (50%) of the minus 400
mesh material in the filter feed was removed. Reason for desliming was to remove
ultrafine particles that retain more surface water. Ultrafine particle may block the
filtration paths, so preventing water from passing or decrease the capillary radii.
The test data, which are summarized in Table 18, showed that filter cake
moistures as low as 10% could be achieved using 1.5-2.5 kg/mt of Reagent RW when
50% of the minus 400 mesh material was removed from the filter feed. The results also
showed that when all of the minus 400 mesh material was removed, a filter cake with as
low as 1.3% moisture could be obtained using only 0.5 kg/mt of RW as dewatering aid.
Although very impressive, the data also showed that a filter product with single digit
moisture levels (i.e., 5% moisture) could be produced without any dewatering aid when
27 |
Virginia Tech | all of the minus 400 mesh material was removed. These results indicate that desliming
may be a viable option for obtaining very low moistures in the Buchanan plant. However,
depending on the size distribution of the feed coal sample, desliming could result in
unacceptably high losses (>30%) of carbonaceous material that is now reporting to the
clean coal product.
Table 18 - Effect of removing slimes (minus 400 mesh material) on dewatering of Buchanan’s
flotation feed
Moisture Content (%wt)
Reagent Dosage
Completely
(kg/mt) No desliming 50% deslimed,
deslimed,
(600 micron x 0) from 38 micron
from 38 micron
0 18.1 16.08 5.00
0.5 17.6 11.10 1.33
1.5 15.2 10.36 --
2.5 14.9 10.12 1.73
Effect of High Speed Mixing
Two series of laboratory filtration tests were conducted to determine the effects of
agitation intensity on the dewatering performance of the Buchanan filter feed. The first
series of tests were conducted using a laboratory shaker to condition the feed samples.
The shaker was a low-energy conditioner that uses reciprocating motion (similar to wrist-
action shaking) to gently mix slurry contained in a 100-ml glass conditioning flask. A
second series of tests were conducted using a 100-ml Plexiglas cell equipped with a
three-blade propeller-type mixer at 1000 rpm. The rotary mixer provided an intense
agitation that is necessary for high-energy conditioning. The feed slurry was conditioned
for 5 minutes in both series of tests.
As shown in Table 19, the moisture reduction was substantially improved when
high-energy conditioning was used. For example, the use of 2.5 kg/t of dewatering aid
reduced the filter cake moisture from a baseline value of 18.2% (no reagent added) down
28 |
Virginia Tech | to 14.6% when using low-energy agitation. The cake moisture was further reduced to
11.7% moisture when high-energy agitation was used at the same reagent dosage of 2.5
kg/t. Similar results were obtained using Reagent RW as the dewatering aid. With this
reagent, the final cake moisture improved from 14.9% to 13.4% at 1.0 kg/t of dewatering
aid and from 14.3% to 13.1% at 2.5 kg/t of dewatering aid. These results clearly
demonstrate the importance of proper conditioning when using the novel dewatering
reagents. The results also indicate that the high-intensity conditioning increases the
adsorption density of dewatering reagents; as a result, lower moisture filter cake product
can be obtained.
Table 19 - Effect of mixing intensity on dewatering of Buchanan’s flotation product
Moisture Content (wt%)
Reagent
Reagent RU Reagent RW
Dosage
(kg/mt)
High Energy Low Energy High Energy Low Energy
Mixing (1) Mixing (2) Mixing(1) Mixing(2)
0 18.2 18.2 18.2 18.2
0.5 15.4 17.1 13.5 15.5
1.0 13.2 14.9 13.4 14.9
1.5 11.9 14.6 13.1 14.7
2.5 11.7 14.6 13.1 14.3
(1) The slurry was agitated at high speed (1000 rpm) using a propeller-type mixer for 5 minutes.
(2) The slurry was agitated at low speed using a laboratory wrist-action shaker for 5 minutes.
29 |
Virginia Tech | Effect of Cake Thickness
The data given in Table 20 show that the baseline moisture (without dewatering
aid) was only slightly lower for the thin cake (18.2% moisture) than for the thick cake
(18.5% moisture). However, the effectiveness of the dewatering aid was much more
pronounced when used in conjunction with the thinner cake. According to the Laplace
equation, the filtration rate is adversely proportional to the length of filtration paths. For
example, the addition of 1.5 kg/t of Reagent RU reduced the cake moisture to 11.9% for
the thin cake (8-11 mm). In contrast, the moisture content of the thick cake (18-20 mm)
was reduced to only 16.0% at the same reagent dosage. The filter cakes typically
produced by the industrial disc filters at the Buchanan plant have a thickness of 20-25
mm under normal operating conditions. It is possible that thicker cakes require longer
drying cycle times.
Table 20 - Effect of filter cake thickness on dewatering of Buchanan’s flotation product using
RU
Moisture Content (wt %)
Reagent Dosage
(kg/mt) Thin Cake Thick Cake
(100 ml slurry, 8-11 mm) (200 ml slurry, 18-20 mm)
0 18.2 18.5
0.5 15.4 16.2
1.0 13.2 16.1
1.5 11.9 16.0
2.5 11.7 15.3
A 62.5 mm diameter Buchner funnel was used. The solid content of the sample was 25-26%. Vacuum
was 68 kPa (20 inches Hg). Cake thickness: 8-11mm (thin cake), 18-20 mm (thick cake). Conditioning
time: 5 minutes; drying cycle time: 2 minutes. The volumes of the slurries were 100 ml for thin cake and
200 ml for thick cake. The slurry was agitated using a high-speed mixer for 5 minutes before dewatering
tests.
30 |
Virginia Tech | CONCLUSIONS
In the present work, dewatering aids were tested on the fine coal of Buchanan
coal (Pocahontas seam). The role of the reagents is to increase the hydrophobicity (or
dewettability) of the coal. It is suggested that the increase in hydrophobicity could be
responsible for the reduction of the capillary pressure in a filter cake and should help
reduce the cake moisture. The novel dewatering aids also decreased the surface tension
of water and increased cake porosity (or capillary diameter), both of which could
contribute to the lower cake moisture.
The test results showed that the use of the novel dewatering aids could
substantially reduce the cake moistures by more than 50% as compared to the case of not
using any dewatering aid. These dewatering aids also increased throughput since the very
fine particles coagulated to become large particles due to hydrophobic coagulation. The
results obtained with a Buchner funnel showed that cake moistures could be reduced
substantially, the extent of which depends on the particle size, cake thickness, drying
time, reagent dosage, conditioning time, reagent type, sample aging, water chemistry, etc.
In addition, the effects of various water treatment strategies were evaluated. These
treatments included the pH control and additions of various sequestrating agents such as
sodium carbonate (Na CO ), ethylenediamine tetraacetic acid (EDTA), sodium silicate
2 3
(Na SiO ), oxalic acid, succinic acid, ammonium oxalate, Na-hexametaphoshate, calcium
2 3
oxide and hydrogen peroxide (H O ).
2 2
The results showed that one way of eliminating Ca2+ ions present in the Buchanan
plant water would be to introduce CO -2 ions. Dewatering tests were conducted using 0.5
3
kg/mt of soda ash in conjunction with Reagents RW and RU. In general, sodium
carbonate was not particularly effective when Reagent RW was used as the dewatering
aid. In fact, the moisture values actually increased slightly when soda ash was added. On
the other hand, the addition of soda ash improved the performance of Reagent RU at
31 |
Virginia Tech | dosages of 1.5 kg/mt or higher. The filter cake moisture was reduced to a very low value
of 12.3% of 2.5 kg/t of RU in the presence of soda ash.
Hydrogen peroxide was used as an oxidizing agent, the test results indicated that
very little improvement in filter cake moisture (only about one percentage point) was
obtained. Similar behavior was observed when EDTA and sodium hexametaphosphate
were used. On the other hand, the addition of sodium silicate generally improved the
performance of the dewatering aids up to a dosage level of 3 g/t of sodium silicate. For
example, the addition of 3 g/t of sodium silicate improved the performance of Reagents
RW and RU by 0.8 percentage points (i.e., 14.3% vs. 13.5% moisture) and 2.0 percentage
points (15.2% vs. 13.2% moisture), respectively. However, further addition of sodium
silicate above 3 kg/t caused an increase in moisture content. The results indicate that the
use of sodium silicate helps in moisture reductions beyond what can be achieved using
dewatering reagents alone.
The moisture content of filter cake could be reduced from 17.3% to 13.3-13.4%
by using 1-5 g/t oxalic acid along with 2.5 kg/t of Reagent RW. Likewise, a filter cake
with as little as 13.3-13.4% moisture could be obtained by using 5-10 g/t oxalic acid
along with 2.5 kg/t RU as the dewatering aid. Similar moisture reductions were obtained
when calcium oxide, ammonium oxalate and succinic acid were used. For example, the
addition of 10 g/t of succinic acid improved the performance of Reagent RW by 2.8
percentage points (i.e., 15.3% vs. 12.5% moisture). Ammonium oxalate improved the
performance of Reagent RW by 1.9 percentage points.
Without a dewatering aid, the lowest moisture values were achieved around
neutral or slightly acidic pH (from 6 to 8). The baseline moisture contents were increased
for both the acidic (pH=4) and alkaline (pH=11) conditions when no reagent was added.
On the other hand, cake moistures in the 13.6-13.9% range were obtained for all pH
values tested except for pH 8 (which gave filter cake moisture of nearly 15%). This
relatively high filter cake moisture corresponded to the pH at which the lowest baseline
moisture was obtained, i.e., when no reagent was employed. These results suggest that a
slightly basic (pH 8) solution may adversely affect the dewatering aids. Additional tests
are needed to verify this observation.
32 |
Virginia Tech | Laboratory filtration tests were conducted to determine the effects of agitation
intensity on the dewatering performance of the Buchanan filter feed. The first series of
tests were conducted using a laboratory shaker to condition the feed samples. The shaker
was a low-energy conditioner that uses reciprocating motion (similar to wrist-action
shaking) to gently mix slurry contained in a 100-ml glass conditioning flask. A second
series of tests were conducted using a 100-ml Plexiglas cell equipped with a three-blade
propeller-type mixer at 1000 rpm. The rotary mixer provided an intense agitation that
may be necessary for high-energy conditioning. The feed slurry was conditioned for 5
minutes in both series of tests. The moisture reduction was substantially improved when
using high-energy conditioning. For example, the use of 2.5 kg/t of dewatering aid
reduced the filter cake moisture from a baseline value of 18.2% (no reagent added) down
to 14.6% when using low-energy agitation. The cake moisture was further reduced to
11.7% moisture when using high-energy agitation at the same reagent dosage of 2.5 kg/t.
Similar results were obtained using Reagent RW as the dewatering aid. With this reagent,
the final cake moisture improved from 14.9% to 13.4% at 1.0 kg/t of dewatering aid and
from 14.3% to 13.1% at 2.5 kg/t of dewatering aid. These results clearly demonstrate the
importance of proper conditioning when using the novel dewatering reagents.
The dewatering tests showed that 40% to 60% moisture reductions could be
achieved from these samples since the hydrophobicity of the particles were increased in
the presence of the dewatering aids. The test results have confirmed that the kinetics of
mechanical dewatering was substantially increased in the presence of the reagents and
increased the throughput of dewatering devices. The data obtained using sequestrating
reagents along with the novel dewatering chemicals developed in Virginia tech show that
the use of these sequestrating reagents is a promising approach for overcoming the water
chemistry problem. The results indicate that the use of sodium silicate help reduce
moistures beyond what can be achieved using dewatering reagents alone.
33 |
Virginia Tech | ABSTRACT
In the absence of fine coal dewatering circuit, the majority of coal preparation
plants use screen bowl-centrifuge to dewater 1mm x 150 micron size clean coal. But the
cost of the dewatering fines sharply increases when the particle is below 150 microns.
For this reason it is often more economical to discard this fine size fraction if it is only a
small part (5 to 10%) of the product stream. The availability of efficient processes that
can reduce the moisture of fine particles will greatly benefit companies if these processes
can be used to reduce the moisture content of fine coal without compromising product
quality.
To investigate the suitability of fine particle dewatering, the novel dewatering
aids, namely RU and RW were tested. Pilot plant tests were conducted on fine coal
samples in order to compare the bench scale dewatering test and the pilot plant test
results.
The tests were carried out using continuous vacuum disc filter. Dewatering results
obtained using the pilot-scale unit are similar with the laboratory dewatering results as
obtained on the Buchanan plant flotation product and US Steel Mining Company’s
Pinnacle Plant’s vibracore composite sample taken from the smith Branch impoundment,
thickener underflow and thickener feed. The use of the novel dewatering aids could
substantially reduce the cake moistures up to 50 % from the fine coal. The dewatering
kinetics were also improved 4 to 6 times in the presence of dewatering aids
The effect of filter disc speed and vacuum pressure were also investigated in the
pilot-scale tests. In the presence of dewatering aids, by simply increasing the filter speed
form 3min/rev to 1min/rev, it is possible to increase the filter capacity by 28% without
adversely impacting the moisture content of the filter cake.
36 |
Virginia Tech | INTRODUCTION
Coal preparation plants consist of simple crushing, sizing and cleaning operations
for coarse and intermediate coal size fractions but commonly involve complex cleaning
circuits for fine size fractions which also require dewatering processes. These can be
broadly classified into three groups; sedimentation, filtration and thermal drying. Since
the market demand is increasing for low-sulfur, high grade coal as a result of economic
and environmental considerations it causes processes to be extended towards the ultra
fine range which accompanies an increase in the yield of ultra fine tailings. Ultra fine
tailings are particularly difficult to dewater because of their large surface area. [1-7]
In the absence of a fine coal dewatering circuit, majority of coal preparation
plants use screen bowl-centrifuge to dewater 1mm x 150 micron size clean coal. The cost
of dewatering fine particles sharply increases when the particle is below 150 microns. For
this reason it is often more economical to discard this fine size fraction of the plant’s run
of mine coal without any attempt to recover the clean coal content of this size fraction if
it is only a small fraction (5 to 10%) of the product stream. The dumping of tailings in the
ponds is not very appropriate nowadays and in the future will be possible only in
exceptional cases because of the lack of suitable sites and also for safety and
environmental considerations.[4, 5, 8-11]
The coal preparation plants, which clean fine coal, utilize filtration technologies
that typically include vacuum disc/drum filters, plate and frame filter presses and
horizontal belt filter for dewatering. When efficiency of dewatering becomes so low,
thermal drying may be required to further dewater fine product.[3, 8, 9]
To avoid the expensive and environmentally important thermal drying, efforts are
being made in the coal preparation sector to counteract this trend both by improvement
and further development of already known dewatering methods and equipment and
testing those not previously used in coal preparation. [3, 8-10]
37 |
Virginia Tech | The availability of efficient processes that can reduce the moisture of fine
particles will greatly benefit companies if these processes can be applied to reduce the
moisture content of fine coal.[8, 9]
To address the problem based on the fine particle dewatering, the novel dewatering aids,
namely RU and RW were tested. Pilot plant tests were conducted on fine coal samples in
order to compare the bench scale dewatering test and the pilot plant dewatering test
results. The test results showed that the use of the novel dewatering aids could
substantially reduce the cake moistures up to half of the total moisture values from the
fine coal. The role of the reagents is to increase the hydrophobicity (or dewettability) of
the coal. In the presence of these reagents, the contact angles of the coal samples
increased preventing the surface of the particles to be wetted by water molecules. It is
suggested that the increase in contact angle could be responsible for the reduction of the
capillary pressure in a filter cake and should help reduce the cake moisture. The novel
dewatering aids also decreased the surface tension of water and increased cake porosity
(or capillary diameter), both of which could contribute to the lower cake moistures.[3]
38 |
Virginia Tech | EXPERIMENTAL
The major objectives of these tests were comparison of pilot scale dewatering test
results with the bench scale dewatering test results and confirm that the newly developed
dewatering aids could be used in the pilot plant tests and through industrial applications.
Reagents RW and RU were used as the dewatering reagents in the pilot scale dewatering
tests. Each of the reagents was tested over a range of dosages that typically varied from 0
to 2.5 kg/ton (0 to 5lb/ton) of the dry coal. Samples were collected around the modules
and analyzed for slurry flow rates and percent solids or moisture content.
Various samples were tested for dewatering tests taken from Consolidation Coal
Corporation’s Buchanan Preparation Plant in Mavisdale, VA and US Steel Mining
Company’s Pinnacle Plant Site near Pineville, WV.
In the tests, the Conditioner Module and Filter Module were required since the feed
slurry for the tests was supplied directly from the Consolidation Coal Corporation’s
Buchanan Preparation Plant in Mavisdale, Virginia. The pilot-scale disc filter tests were
conducted on flotation product samples.
The Column Flotation, Conditioner and Filter Modules were set up in the Virginia
Tech’s Plantation Road Pilot Plant Facility in Blacksburg and used to test three of the
coal samples taken from the US Steel Mining Company’s Pinnacle Plant Site near
Pineville, WV. The samples were vibracore composite sample taken from the Smith
Branch Impoundment, a slip stream sample of current thickener feed and thickener
underflow.
The schematic diagram of the Column Module shown in Figure 2 consists of a 30-
cm diameter by 3-m tall flotation column. The column is equipped with the Microcel
sparging system that circulates a portion of the slurry from the bottom of the column
through an in-line static mixer. Up to 100 liters/minute of air is supplied to the static
mixer by a rotary air compressor. Coal slurry is fed to the column from an agitated tank
using a variable-speed centrifugal pump. Pulp level in the column is maintained by
adjusting the tailings flow rate using a pneumatic control valve. The valve actuates based
39 |
Virginia Tech | on readings from a pressure transducer mounted in the side of the column. Wash water is
added to the froth to minimize the entrainment of fine mineral matter. Chemical metering
(reagent) pumps are used to add the desired dosages of frother and/or collector to the feed
slurry.
Column Module
12
Dimensions
10 in feet.
8
6
4
Front Side
2
View View
0
00 22 44 66 88 00 22 44 66 88
Figure 2. Schematic diagram of the Column Module
The pilot scale flotation column was capable of producing 22-34 kg/hr of
concentrate grading 7-8% ash from a feed containing (on average) 42 % ash at recoveries
of about 95%, and that the filter produced from 8 to 35 kg/hr of cake (depending on
reagent dosage) from feed streams containing around 7-8% solids.
A schematic diagram of the Conditioner Module is shown in Figure 3. The
module incorporates two 20-liter conditioning tanks that are operated in series to provide
up to 10 minutes of conditioning time. The conditioning tanks are equipped with single-
impeller mixers that can be varied in speed from 0 to 2500 rpm using electronic
controllers. To ensure that coarser particles do not settle when low feed rates are used,
40 |
Virginia Tech | the slurry in the conditioning tanks is continuously circulated through a head tank using a
centrifugal pump. The head tank is equipped with an automated sampling system that
consists of an electronic timer and a pneumatic sample cutter. During operation, the
sampling system is used to divert a defined portion of circulated slurry to the Filter Module
(or any other downstream operations). A chemical metering (reagent) pump is used to add
the proper dosage of dewatering aid to the feed slurry as it enters the conditioning tanks. To
obtain a consistent feed rate, a small peristaltic pump was installed at the plant to pump feed
slurry from the plant’s filter feed box to the conditioning module.
Conditioner Module
12
Dimensions
10 in feet.
8
6
4
Front Side
2
View View
0
00 22 44 66 88 00 22 44 66 88
Figure 3. Schematic diagram of the conditionar module
The tests were conducted with a Pilot-Scale Disc Filter that was manufactured by
Peterson Filter Company, Salt Lake City, Utah (Schematic diagram 4). The principle of
construction of disc vacuum filter is a disc mounted on a horizontal shaft, having
interchangeable elements which can be changed for fitting and removing filter cloths. It
41 |
Virginia Tech | • Dual filtrate sumps: 25 gal capacity each
• Connected HP: 2.25
• Dimensions: 5ft. High x 5ft wide x 4ft deep.
• Weight: 1,800 lb.
In some cases, the use of the dewatering aids increased the rate of dewatering
kinetics which results in a substantial increase in cake thickness, requiring a longer
drying time to achieve desired cake moisture. To minimize this problem, the rotational
speed of the disc filter can be increased, which will in turn shorten the drying cycle time.
This may be referred to as a ‘dilemma’ faced in using disc filters for fast-dewatering
materials. A solution to this problem was splitting a vacuum manifold into two separate
lines; one directed to the submerged filter sectors and the other to those that are open to
the air. The pilot-scale disc was modified to incorporate this “dual vacuum system,” as
shown in Figure 5. It is equipped with a pressure reducer, which allows the vacuum
pressure for the bottom sectors of the disc to be reduced, while the upper sectors to have
full vacuum pressure.
Gas
Compressor
Vacuum
Pump
Pressure
Reducer
Mechanical
Controller
High-Vacuum
Low-Vacuum
Figure 5 Figure illustrating the principle of the dual vacuum system
This property makes it possible to control the vacuum pressures during the cake
formation (or pick-up) time and drying cycle time independently from each other. When
43 |
Virginia Tech | of the cake were reduced from 17.6% to 16.1% and 14.4%, respectively. Similar results
were obtained with RW as a dewatering aid. These values correspond to a 15-20%
moisture reduction in the filter product.
Table 22 - Effect of reagent addition on laboratory scale dewatering of Buchanan’s flotation
product
Reagent Dosage Moisture Content (%wt)
(kg/mt Reagent W Reagent U
0 17.6 17.6
0.5 16.2 16.1
1.5 16.2 14.9
2.5 15.0 14.4
Dewatering results obtained using the pilot-scale unit agreed well with the
laboratory dewatering results as obtained on the Buchanan plant flotation product. As
shown in Tables 21 and 22, the moisture contents of the filter products for both the
laboratory and the pilot-scale test programs are 14.5% at approximately 2.5 kg/mt (5
lb/ton) of Reagent U.
The effect of filter disc speed was also investigated in the pilot-scale tests. Table 23
summarizes the results obtained by increasing the filter disc speed from 3 to 1 min/rev for
dewatering of the plant flotation product. The product rate increased from 83.3 to 116.6
kg/hr, with little change in the moisture content of the filter cake. The results show that it
would be possible to increase the filter capacity by 28% without adversely impacting the
moisture content of the filter cake. Tests were conducted on Buchanan plant flotation
product without dewatering reagents.
46 |
Virginia Tech | Table 23 - Effect of pilot-scale filter disc speed on filter cake production rates and moisture
content
Filter Disc Speed Product Rate Moisture Content
(mins/rev) (kg/h dry) (% wt)
3.0 83.3 16.9
2.0 108.6 16.7
1.0 116.6 17.3
A series of pilot-scale dewatering tests were conducted to study the effects of
different vacuum levels on moisture reduction. The results are presented in Table 24. The
data show that it is possible to reduce the product moisture from 19.7% to 16.0% by
simply increasing the vacuum level from 5 to 15 inches Hg. It seems that a further
increase in vacuum from 15 to 20 inches Hg is not advantageous in terms of further
lowering the moisture contents of the filter products. Tests were conducted on Buchanan
plant flotation product without dewatering reagents
Table 24 - Effect of pilot-scale filter vacuum level on filter cake moisture contents
Vacuum Moisture Content
(kPa) (% wt)
34 19.7
51 16.7
68 17.1
It should be mentioned here that the disc filters in the Buchanan preparation plant
are currently operated at vacuum levels of only 10.5-11.0 inches Hg. Because of such low
vacuum levels, the plant filter product typically contains 21-22% moisture. The present
work shows that by increasing the vacuum levels from 36 to 51-54 kPa, the plant could
probably obtain a filter product with 17-18% moisture. Besides dewatering aid addition,
vacuum level is one of the important operating conditions determining the final product
moisture in the filter cake.
47 |
Virginia Tech | Table 25 gives the pilot scale test results obtained on the Buchanan plant flotation
product using RW as the dewatering aid at various vacuum pressures. As shown, the
moisture content in the filter cake was reduced with increasing vacuum pressure. At
vacuum pressure increasing from 37kpa to 68kpa, the moisture contents of the cake were
reduced from 18.2 to 16.8% and 16.3%, respectively.
Table 25 - Effect of pilot-scale filter vacuum level on filter cake moisture contents.
Reagent Vacuum Moisture Content
(inHg) (% wt)
RW 37 18.2
(2kg/ton) 54 16.8
68 16.3
The test results, which are given in Table 26, indicate that higher improvement in
filter cake moisture (about two percentage point) was obtained when using RU as
dewatering aid. As the vacuum levels increased from 37 to 68kpa, the moisture content of
the cake were reduced from 16.8 % to 14.5% and 14.3%, respectively.
Table 26 - Effect of pilot-scale filter vacuum level on filter cake moisture contents.
Reagent Vacuum Moisture Content
(inHg) (% wt)
RU 37 16.8
(2kg/ton) 54 14.5
68 14.3
48 |
Virginia Tech | 2. Dewatering Test Results on the US Steel Mining Company’s Pinnacle Plant’s
Coal Samples
Table 27 gives the laboratory test results obtained on the Pinnacle pond sample
using RW as the dewatering aid. At 0.5 and 2.5 kg/mt RW additions, the moisture
contents of the cake were reduced from 28.7% to 24.6% and 14.3%, respectively. The
lower value corresponds to a 50% moisture reduction in the filter product.
Table 27 - Effect of Reagent addition on dewatering of Pinnacle pond sample
Reagent Dosage Moisture Content (%wt)
(kg/mt) Reagent W
0 28.7
0.5 24.6
1.0 22.9
1.5 22.5
2.5 14.4
Similar tests were conducted on the thickener underflow and thickener feed
samples using RW and RU as dewatering aids. Results for the thickener underflow
sample are given in Table 28, which shows the moisture content of the filter product
decreases with increasing RW and RU additions from 0.5 to 2.5 kg/mt. The RW addition
rate of 2.5 kg/mt reduced cake moisture contents from 31.4% to 22.1%, giving a
percentage moisture reduction of about 30%.
Table 28 - Effect of reagent addition on dewatering of Pinnacle thickener underflow sample
Regent Dosage Moisture Content (%wt)
(kg/mt) Reagent W Reagent U
0 31.4 31.4
0.5 30.3 28.8
1.0 28.0 28.2
1.5 22.7 --
2.5 22.1 27.8
49 |
Virginia Tech | Table 29 gives the results for the thickener feed sample. In this case, RW and RU
gave similar results - cake moisture contents down from 29.5% to 21.1% and 21.8%,
respectively, and percentage moisture reductions of about 28 and 26% at addition rates of
2.5 kg/mt.
Table 29 - Effect of reagent addition on dewatering of Pinnacle thickener feed sample.
Regent Dosage Moisture Content (%wt)
(kg/mt) Reagent W Reagent U
0 29.5 29.5
0.5 26.5 27.3
1.0 24.3 25.2
1.5 22.2 23.1
2.5 21.1 21.8
Reagent W and U were used as dewatering aids in pilot scale dewatering tests.
Table 30 gives the pilot scale dewatering results obtained on the Smith Branch
Impoundment sample. These results show that the moisture contents of the filter cakes
are substantially decreased in the presence of dewatering aid, at reagent dosage rates of
1.0-2.5 kg/mt of RW, filter cake moisture content was reduced from 28.4% to 17.7-
16.3%. Cake thicknesses were as high as 16 mm when using RW as the dewatering aid.
Even at this cake thickness, moisture reductions were significant. The percentage
moisture reduction was 43%. Filter effluents were also much cleaner when dewatering
aids were used, indicating that filter recoveries increased substantially in the presence of
dewatering aid.
50 |
Virginia Tech | Table 30 - Effect of RW addition on the pilot scale dewatering of the Pinnacle-Smith Branch
Impoundment sample.
Regent Dosage Moisture Content (%wt)
(kg/mt) Reagent W
0 28.4
0.5 19.6
1.0 17.7
1.5 17.2
2.5 16.3
Table 31 and 32 give the pilot scale dewatering results obtained on the Pinnacle
coal thickener feed sample. In this series of tests, RW was used as the dewatering aid.
As shown in Table 31, the moisture content of the filter cake is reduced from 29.4% to
21.5% by the addition of 2 kg/mt RW.
Table 31 - Effect of RW addition on the pilot scale dewatering of the Pinnacle-thickener feed
sample.
Regent Dosage Moisture Content (%wt)
(kg/mt) Reagent W
0 29.4
0.5 --
1.0 --
2.0 21.4
2.5 --
Of particular note in these two mobile plant tests is that the filtration performance
of both the pilot-scale filter and Reagent W are essentially identical to that achieved in
the laboratory.
The test results given in Table 32 show that the moisture content of the filter cakes
are significantly decreased when RW and RU are used as dewatering aids. For example,
filter cake moisture contents were reduced from 28.0% to 23.4% at 0.95 kg/mt and 20.3%
at a 3.7 kg/mt reagent dosage. As shown, RU performs as well as RW in terms of
moisture reduction. The moisture content in the filter cake was reduced from 28% to
51 |
Virginia Tech | CONCLUSIONS
The objective of the dewatering tests at the pilot plant was to verify some of the
results obtained from the laboratory-scale batch dewatering tests. The verification tests
were carried out using continuous disc filter. Dewatering results obtained using the pilot-
scale unit are similar with the laboratory dewatering results as obtained on the Buchanan
plant flotation product. The test results showed that in some cases the moisture content of
the pilot plant filters were 5 to 10% higher than laboratory dewatering tests. However, the
final moisture reduction of both filters was closer to each other.
The effect of filter disc speed was also investigated in the pilot-scale tests. By
increasing the filter disc speed from 3 to 1 min/rev for dewatering of the plant flotation
product, the product rate increased from 83.3 to 116.6 kg/hr, with little change in the
moisture content of the filter cake. It is possible to increase the filter capacity by 28%
without adversely impacting the moisture content of the filter cake.
Some pilot.-scale dewatering tests conducted to study the effects of different
vacuum levels on moisture reduction show that it is possible to reduce the product
moisture from 19.7% to 16.0% by simply increasing the vacuum level from 34 kPa to 51
kPa.
The bench scale dewatering test results obtained on the Pinnacle pond sample
using RW as the dewatering aid at showed that the moisture content in the filter cake was
reduced with increasing reagent addition. At 0.5 and 2.5 kg/mt RW additions, the
moisture contents of the cake were reduced from 28.7% to 24.6% and 14.3%,
respectively. The lower value corresponds to a 50% moisture reduction in the filter
product.
Similar bench scale dewatering tests were conducted on the thickener underflow
and thickener feed samples using RW and RU as dewatering aids. The RW addition rate
of 2.5 kg/mt reduced cake moisture contents from 31.4% to 22.1%, giving a percentage
moisture reduction of about 30%. This is less than the case of the impoundment sample,
and can attributed to the different mean particle sizes of these two samples. Mean
53 |
Virginia Tech | particle size was approximately 32 microns for the impoundment sample and 15 micron
for the thickener underflow sample.
RW and RU gave similar results - cake moisture contents down from 29.5% to
21.1% and 21.8%, respectively, and percentage moisture reductions of about 28 and 26%
at addition rates of 2.5 kg/mt for the thickener feed sample.
The pilot scale dewatering results obtained on the Smith Branch Impoundment
sample show that the moisture contents of the filter cakes are substantially decreased in
the presence of dewatering aid, at reagent dosage rates of 1.0-2.5 kg/mt of RW, filter
cake moisture content was reduced from 28.4% to 17.7-16.3%. Cake thicknesses were as
high as 16 mm when using RW as the dewatering aid (vs. 3-6 mm without). Even at this
cake thickness, moisture reductions are significant. Filter effluents were also much
cleaner when dewatering aids were used, indicating that filter recoveries increased
substantially in the presence of dewatering aid.
The pilot scale dewatering results obtained on the Pinnacle coal thickener feed
sample shows that the moisture content of the filter cake is reduced from 29.4% to 21.5%
by the addition of 2 kg/mt RW. The test results again show that the moisture content of
the filter cakes are significantly decreased when RW and RU are used as dewatering aids.
Filter cake moisture contents were reduced from 28.0% to 23.4% at 0.95 kg/mt and
20.3% at a 3.7 kg/mt reagent dosage. As shown, RU performs as well as RW in terms of
moisture reduction. The moisture content in the filter cake was reduced from 28% to
20.6% in the presence of 3.1 kg/mt RU.
These results are in good agreement with those obtained in the laboratory test
work and the pilot-scale filter and Reagent W are essentially identical to that achieved in
the laboratory.
REFERENCES
1. Apling, A., Mineral processing technology: B. A. Wills, Pergamon Press, Oxford
(1992), ISBN 0080418856 ISBN 0080418724 [UK pound. Corrosion Science,
1994. 36(4): p. 743-744.
2. Abbott, J., Advanced coal preparation monograph series volume VII part 16
control concepts: C.J. Clarkson Australian Coal Preparation Society,
54 |
Virginia Tech | Site Characterization, Sustainability Evaluation and Life Cycle
Emissions Assessment of Underground Coal Gasification
Zeshan Hyder
Abstract
Underground Coal Gasification (UCG), although not a new concept, is now attracting
considerable global attention as a viable process to provide a “clean” and economic fuel
from coal. Climate change legislation and the declining position of coal reserves (i.e., deeper
and thinner seams) in many parts of the world are promoting and fueling the UCG
renaissance.
This research presents an analysis of operational parameters of UCG technology to
determine their significance and to evaluate the effective range of values for proper control
of the process. The study indicates that cavity pressures, gas and water flow rates,
development of linkage between wells, and continuous monitoring are the most important
operating parameters.
A protocol for the selection of suitable sites for UCG projects is presented in this study. The
site selection criteria are developed based on successes and failures of previous
experiments and pilot studies. The criteria take into account the site characteristics, coal
quality parameters, hydrology of the area, availability of infrastructure and regulatory and
environmental restrictions on sites. These criteria highlight the merits and demerits of the
selected parameters, their importance in site selection and their economic and
environmental potentials.
Based on the site selection criteria, a GIS model is developed to assist in selecting suitable
sites for gasification in any given area of interest. This GIS model can be used as a decision
support tool as well since it helps in establishing the tradeoff levels between factors,
ranking and scaling of factors, and, most importantly, evaluating inherent risks associated
with each decision set. |
Virginia Tech | Abstract
The potential of UCG to conform to different frameworks defined to assess the capability
and potential of any project that merits the label, “sustainable,” has been evaluated. It has
been established that UCG can integrate economic activity with ecosystem integrity,
respect for the rights of future generations to the use of resources and the attainment of
sustainable and equitable social and economic benefits. The important aspects of UCG that
need to be considered for its sustainable development are highlighted.
In addition, the environmental benefits of UCG have been evaluated in terms of its potential
for reduction in greenhouse gas (GHG) emissions. The findings indicate that UCG
significantly reduces GHG emissions compared to other competitive coal exploiting
technologies. A model to compute the life cycle greenhouse emissions of UCG has been
developed, and it reveals that UCG has distinctive advantages in terms of GHG emissions
over other technologies and competes favorably with the latest power generation
technologies. In addition to GHG emissions, the environmental impacts of these
technologies based on various impact assessment indicators are assessed to determine the
position of UCG in the technology mix. It is clear from the analysis that UCG has prominent
environmental advantages and has the potential to develop and utilize coal resources in an
environmentally friendly and economically sound manner.
iii |
Virginia Tech | Chapter 1 - Introduction
1.1 Introduction
Underground or insitu coal gasification (UCG) is a process involving insitu burning and
conversion of coal into gaseous product, commonly known as synthetic gas or syngas. This
syngas is at high temperature and pressure and can be used for several processes including
but not limited to electricity, heat and power generation, feedstock for several chemical
processes such as Fischer-Tropsch, synthetic natural gas & hydrogen production, iron
reduction and chemical products like ethylene, polyolefin, methanol, petrol, acetic acid and
formaldehyde [Liu, Mallet et al. 2003; Courtney 2009; Oukaci 2009]. In its simplest form,
the process involves an injection and a production well, drilled up to the coal seam. Then
steam, air or oxygen is injected under pressure through the injection well to start
combustion process and product gases at high temperature and pressure are collected at
the production well.
2 4 2
The composition of syngas is largely H , CH , CO and CO , with a calorific value of 850 -1200
3 3
kcal/Nm (~ 3.5 – 5 MJ/m ) [Ghose and Paul 2007]. However, the composition and calorific
value of syngas varies according to the site characteristics (coal ranks, strata type, depth
and amount of moisture etc.) and type of gases used for injection (air, oxygen or steam).
3
The calorific value for air injected syngas typically ranges from 4.0-5.5 MJ/m and
approximately doubles when oxygen is injected instead of air [Walker 1999].
1.2 Problem Statement
Underground or Insitu Coal Gasification (UCG or UICG), though not a new concept, is now
attracting considerable global attention as a viable process to provide a “clean” and
economic fuel from coal. This technology has the potential to exploit energy from low
grade, deep seated, thin coal seams in an economic, environment friendly and sustainable
manner. This technology can be applied to abandoned coal mines, remnants of already
exploited reserves and deposits considered uneconomic and technically difficult for
1 |
Virginia Tech | Chapter 1-Introduction
conventional mining methods. Despite these economic, environmental and technical
benefits, this technology has the potential to create hazards of atmospheric emissions,
ground water contamination, uncontrollable cavity growth & underground fires,
2
subsidence, CO management and human impacts like noise, dust, and increased traffic, if
the process is not properly managed. Thus, in order to manipulate the full economic and
environmental potential of this technology, it is imperative that its development is in line
with sustainable development principles, based on proper site selection criteria and
environmental impact assessments. As this is a developing technology, many research gaps
need to be filled. Some of the most prominent aspects dealt in this research include
development of site selection criteria, GIS modeling of selected sites, Life cycle assessment
to compare environmental impacts with other competitive technologies and sustainability
assessment of UCG.
1.3 Research Objectives
Since UCG is a technology that is currently in development stage, many aspects require
focused research to develop it into a mature and commercially acceptable alternate for
exploitation of energy from coal. The focus of this research is on the following core
ob jectives:
1. Operational Parameters of UCG
a. Review of important operational parameters
b. Analysis of developments in operational criteria
2. Development of site selection criteria
a. Outline important site selection parameters for UCG
b. Evaluate the importance of defined parameters for site selection and resource
assessment
c. Application of this criteria to identify potential UCG sites
d. Development of GIS Model based on defined parameters/criteria
3. Sustainability assessment of UCG operations
a. Assessment of UCG within sustainable development frameworks
b. Identify important principles essential for sustainable development of UCG
2 |
Virginia Tech | Chapter 2 - Literature Review
2.1 History of UCG
Underground coal gasification (UCG) is not a new concept and can be traced back to as
early as 1868, when Sir William Siemens, a German scientist first presented the idea of
burning ash-rich coal underground before the Fellows of the Chemical Society of London
[Olness and Gregg 1977]. The famous Russian scientist, Dmitry Mendeleyev mentioned the
same idea some twenty years later in 1888, when he remarked that in the future coal
would be burned underground and supplied as gas via pipes for consumption. Mendeleyev
exclusively suggested his depiction of UCG during a visit to Kizel plant in Ural Russia in
1899, where he gave the idea about controlling the underground coal fires by drilling of
injection and production well, with supply of limited air quantity through one pipe and
syngas withdrawing by suction though other pipe [Klimenko 2009]. In 1909, Great Britain
issued first patent for UCG to an American, A.G. Betts, who presented two different methods
of gasifying seams too thin or to poor for mining [Olness and Gregg 1977]. Around 1910,
UCG was included in the proposals suggested by a special committee of Britain’s leading
scientists investigating national energy sources. An English chemist, Sir William Ramsey,
who was also serving on the committee, took great interest in the idea of UCG and
expanded Bett‘s idea. In 1912, he presented his idea and its merits at the International
Smoke Abatement Exhibition, London. His method involved a borehole in the coal and
passing of air and steam through the hole to initiate water gas reaction. This first trial
experiment got funding but never occurred because of Ramsey’s death and the outbreak of
World War I[Olness and Gregg 1977] .
For the next several years, no further progress was reported in underground gasification.
However, the idea of gasifying coal underground and Ramsey’s reports attracted Lenin,
who was living in exile in Zurich at that time. Lenin liked the idea of UCG as it had potential
of reducing labor requirements and improved working conditions for miners, freeing them
from hard underground labor. This reduction of labor could contribute to his idea of
6 |
Virginia Tech | Chapter 2 - Literature Review
shortening the working day to 7 hours or less, a major concern of socialist society. On May
4 1913, Lenin published an article in Pravda highlighting the potential of UCG [Olness and
Gregg 1977]. This brought the UCG concept into the political lime light and placed it in the
public. Based on the same potentials envisaged by Lenin, Joseph Stalin championed the UCG
program in Former Soviet Union. First trials started in about 1928 and for the next five or
so decades, Soviets ruled in UCG technology, with total Soviet efforts far exceeding the
combined efforts of other nations and commercial UCG plant still working in Angren,
Uzbekistan [Brown 2012].
The energy shortage between 1944 and 1959 promoted interest for UCG in Western
European coal mining countries and Czechoslovakia, France, Italy and Poland conducted
UCG tests. In Britain, experimental work started in 1949 at Newman Spinney in Derbyshire
and in Bayton Worcestershire the following year [Thompson, Mann et al. 1976]. However,
the work on UCG stopped in the 1960s due to abundance of energy and lower oil prices.
European work started again in the 1980s and several UCG tests were carried out in
Western Europe [Couch 2009]. The tests were carried out in Belgium in 1982 and in France
during 1983 and 1984. In 1989, the European Working Group recommended a series of
trials to determine commercial feasibility of UCG in thinner and deeper coal seams of
Europe. The first of such trials was undertaken in 1992-1999 by Spain, UK and Belgium
with the support of European Commission and was known as “the Spanish trial”. For this
trial two fields were proposed, one having coal at a depth of about 600 m to represent
southern Europe coal fields and the second at 900 m in bituminous coal typical of mid and
northern European fields [Creedy, Garner et al. 2001]. This project demonstrated key
technological and research requirements for further development of this technology and
highlighted likely environmental impacts.
In the U.S., research and development about UCG started in the 1940s. The program
accelerated in the 1970s due to energy crisis and more than 30 experiments were
conducted between 1972 and 1989 [PWC. 2008]. The U.S. Department of Energy funded
most of these trials [Klimenko 2009]. Most notable of these trials were Hoe Creek,
Wyoming and Centralia, Washington. These tests were well documented and provided the
basic information for several pilots and commercial tests in other countries. However,
7 |
Virginia Tech | Chapter 2 - Literature Review
2.3 Historical Developments
In early 1930s, the Russians started an extensive research program about UCG. During the
next 50 years, they conducted several pilot and commercial trials on UCG. In earlier trials
they tried to burn the coal adjacent to the road ways and drifts, however, this process did
not prove successful as it burned the produced gas and allowed the air to bypass the
burning zone, so they tried boreholes instead of roadways, initially with the holes drilled
from the roadways but later on from the surface [Olness and Gregg 1977]. At Lisichansk,
the deviated holes were drilled, which was an impressive achievement at that time
[Klimenko 2009]. The Russians also developed a method to gasify steeply dipping coal
seams, with the borehole drilled from the surface, as shown in Figure 2.4. A steeply sloping
seam at Yushno Arbinsk, USSR was developed for UCG, drilling boreholes from the surface
[Thompson, Mann et al. 1976].
Figure 2.4: Drilling into a steeply inclined seam from the surface
10 |
Virginia Tech | Chapter 2 - Literature Review
The first trial at Britain started in 1949 at Newman Spinney, adjacent to an opencast site
and the seam was developed for UCG by drilling a hole from the exposed surface, initially
keeping the hole above the seam and then allowing it to cross through the seam. In another
trial at Bayton England, the pneumatic linkage method was tried successfully, with high-
pressure air used to promote the linkage and enlarge the cavity between boreholes. In
1958, the National Coal Board England, planned a trial to develop the site at Newman
Spinney using the blind borehole technique, where holes were drilled from a roadway or
tunnel into the seam and terminated at blind or dead end [Thompson, Mann et al. 1976].
Figure 2.5 is a schematic of this technique.
Figure 2.5: Blind borehole method, Newman Spinney
At Newman Spinney another approach to gasifier design was tried based on Russian, Polish
and Belgian work at that time in which the entire length of the gallery was ignited by
means of wooden sleepers and a broad reaction front advanced until the final form of voids
was achieved. From 1962-1966, the working site was exposed by opencast mining and
11 |
Virginia Tech | Chapter 2 - Literature Review
provided a visual look of the site which confirmed the shape and size of the cavities to a
remarkable degree [Thompson, Mann et al. 1976], as shown in Figure 2.6.
Figure 2.6: Voids left at termination, Newman Spinney
2.4 Current Developments
Currently the focus of research is to understand the behavior, dynamics and control of
cavity growth, the thermodynamic changes during the process, heat inflow and out flow,
flame propagation, linkage between injection and production well, synergies between UCG
and CCS and burn cavity growth and process modeling [Hobbs, Radulovic et al. 1993;
Blinderman, Saulov et al. 2008; Daggupati, Mandapati et al. 2010; Roddy and Younger
2010; Saulov, Plumb et al. 2010; Stanczyk, Smolinski et al. 2010]. Main technique used for
UCG is Linked Vertical Wells (LVWs), consisting of two wells linked underground using
high-pressure air, water or reverse combustion. Russians developed this technology and it
was used at Chinchilla, Australia recently [Walker, Blinderman et al. 2001]. There are
several variants of this technology. They all are based on difference in techniques to
12 |
Virginia Tech | Chapter 2 - Literature Review
is the super daisy concept that involves drilling blind holes from a shaft. Its layout is a shaft
of 5.0 m in diameter and 300-400 jet stingers [Palarski 2007]. Another development is use
of the CRIP method pioneered in the U.S. by Lawrence Livermore National Laboratory,
during the trials in 1980s. It involves retraction of the burning front successively, thus
exposing unreacted coal surfaces to oxidation and gasification [Clean Coal Limited 2007]. It
has two variants, Linear CRIP and Parallel CRIP. This method is used in the Spanish trial by
Linc Energy Ltd. and by Carbon Energy Pty Limited (CEPL) in Blood wood Creek with
parallel holes [Carbonenrgy Ltd. 2011].
The Russians have designed a new technique for UCG development using in-seam
boreholes and CRIP type operation [Couch 2009]. In the new technique, the in-seam
borehole makes the linkages between the injection wells and the oxygen or air supply point
moves and retracts successively like CRIP, thus improving gasification process.
14 |
Virginia Tech | Chapter 2 - Literature Review
Temperature 200-550 550-900 0C > 900 0C
0C
Oxidation Zone
Reducing Zone
Drying and Pyrolysis
Zone
pyrolysis reduction oxidation
4 2 2 2 2 2
Coal CH + H O C + H O CO + H C + O CO
2 2 2
CO+ CO C + CO 2CO C + ½ O CO
2 2 4 2 2
H + C CO + 2H CH C O+ ½ O CO
2 2 2 2 2 2
CO + H O CO + H Coal + O CO +CO+H O
hydrocarbons
Figure 2.8: Schematic of processes involved in UCG
2.7 Benefits of UCG
UCG has distinct economic and environmental benefits over conventional coal mining,
surface gasification and even coalbed methane drainage. There is a 20 times more energy
yield from a given coal resource, when UCG is applied instead of CBM [Meany and Maynard
2009]. Creedy et al. suggest that as a method of exploiting coal, UCG represents a
substantial environmental improvement on the combination of coal mining and surface
c2o.7a.l1 c omEcbounsotimoni c[ CBreeneedfyi,t sG arner et al. 2001].
UCG has the potential of exploiting coal resources, which are uneconomic for underground
coal mining or are too deep to mine due to technical and safety restrictions. UCG can
enhance workable coal reserves due to its applicability to deposits not minable through
16 |
Virginia Tech | Chapter 2 - Literature Review
conventional mining methods such as coal deposits having low calorific values, deposits
having thin seams or are steeply inclined or having thick seams that are deep seated [Lamb
1977]. That is why Burton et al. suggest a 300-400% increase in the recoverable coal
reserves of the U.S. through application of UCG [Burton, Friedmann et al. 2006]. Although in
the current market, natural gas is more attractive power generation source, in terms of cost
and acceptability when compared to UCG, its reserves are finite and at some time in future,
it will be more economical and strategically beneficial to replace or complement natural
gas with UCG in power generation schemes [Creedy, Garner et al. 2001] .
Another economic benefit of UCG is reduced capital cost due to elimination of mining,
transportation and storage of coal. There is no requirement of waste & ash management
facilities and surface gasifiers. The syngas can be used directly as a fuel and feedstock for
power generation and other chemical industries, after processing through onsite gas
cleaning plant.
Another important advantage of UCG is its applicability to all kind of coals from lignite to
anthracite [Burton, Friedmann et al. 2006] as demonstrated successfully in the surface
gasifiers and also tested in UCG pilots and trials. However, some types of coals are easier to
gasify than other types, for example swelling and coking coals can decrease permeability
and porosity of coal upon heating. Low rank coals are preferred for gasification because
they are generally shallow, easy to ignite and shrink on heating resulting in enhanced
permeability and improved linkages between injection and production wells [Bialecka
2009]. This increases the economic viability of the UCG process as generally low ranked
coals are given downgraded preference in the conventional mining. Though, the calorific
value of syngas produced from lignite is generally in the lower range, several experiments
have successfully resulted in the increased calorific value of syngas through the control of
cavity geometry, burning process and type of oxidant used i.e. air, stream or oxygen
[Hongtao, Feng et al. 2011; Stanczyk, Howaniec et al. 2011] .
UCG can work on any seam thickness, though it is preferable to have a seam thickness of
more than 2.0 m [Burton, Friedmann et al. 2006]. As reported by Shafirovich and Varma,
Ergo Exergy claims that coal seams as thin as 0.5 m may be gasified through UCG
17 |
Virginia Tech | Chapter 2 - Literature Review
[Shafirovich and Varma 2009]. However, thicker coal seams (>2 m) require fewer
boreholes to gasify more amount of coal and provide economic and environmental
advantages [Sury, Kirton et al. 2004].
UCG trials have been conducted on coal seams at varying depths, for example in Russia and
U.S. coal seams as deep as 30 m to 350 m were gasified whereas in Europe, trials at much
deeper depths (600-1200 m) were carried out [Shafirovich and Varma 2009]. At shallower
depths, the problems of gas leakages and water contamination are more pronounced. At
deeper depths, most of the aquifers are either saline or not potable, so the problem of
aquifer contamination is reduced. Drilling at greater depths may increase the operating
cost but new technological developments in drilling have made it possible to operate at
greater depths without any technical and operational difficulties, generally faced in the
past. New technologies and design solutions have promoted development of deep coal
seams, increased control over rock pressures and reduction in the well requirements, thus
decreasing drilling costs and encouraging UCG application to abandoned mines [Zorya, JSC
Gazprom et al. 2009].
Seam inclination or dip is not a restraining factor for UCG site selection [Shafirovich and
Varma 2009] and steeply dipping seams are preferred for UCG as they are considered less
economical and technically difficult to mine through conventional mining methods as
compared to horizontal seams [Lamb 1977]. Another potential economic benefit of UCG is
i2t.s7 s.2yn eErngyv iwroitnhm caernbtoanl Bcaepnteufriets a nd sequestration [Friedmann 2010].
UCG possesses several environmental advantages as well. During the process, ash and
heavy metals remain underground [Fergusson 2009], thus reducing cost and efforts for
waste management. Sulfur and nitrogen come out to the surface with product gas but
conventional sour gas cleaning technologies can remove sulfur from the gas [Fergusson
2009]. As reported by Meany and Maynard, ash content of syngas is one-seventh (1/7th)
x
that of coal burnt on surface and various concentrations of toxic nitrogen oxides (NO ) and
2
sulfur dioxide (SO ) are greatly reduced by UCG [Meany and Maynard 2009].
18 |
Virginia Tech | Chapter 2 - Literature Review
Due to elimination of mining, there is no overburden dump, no spoil dumps and no ash or
tailing dams as most of the waste remain underground [Meany and Maynard 2009]. Creedy
et al. have highlighted following environmental benefits of UCG over coal mining [Creedy,
Garner •e t al. 2001]:
• Reduction in dust, noise and visual impact on surface
• Lower water consumption
• Lower risk of surface water pollution
• Reduced methane emission
• No Dirt handling and disposal at mine
• No coal washing and fines disposal at mine site
• No ash handling and disposal at power station sites
• No coal stocking and transport
• Smaller surface footprint
No mine waste recovery and significant surface hazard liabilities on
abandonment
Similarly, the process creates minimal surface disruption, thus resulting in reduced land
acquisition and rehabilitation requirements [Ghose and Paul 2007]. Another important
2
aspect of UCG is its synergy with carbon capture and sequestration (CCS). CO is the main
component of UCG syngas and may be present in the range of 25-40%. Integration of UCG
and CCS may result in a critical climate change mitigation technology capable of producing
power from coal and many studies suggest it as a low cost, above ground, low carbon form
of coal power production [Redman, Fenerty et al. 2009]. The ideal scenario is to capture
2
CO from the product gas at the source and sequester it near the project site to avoid
2
transportation cost. The technology to capture CO from the syngas and to store it in the
2
geological formations is readily available; however, research is in progress to store CO in
the voids and cavities created by UCG processes [Friedmann 2009; Roddy and Younger
2010].
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An important environmental concern about UCG is its impact on groundwater. It can
contaminate nearby aquifers through gas leakage and leachate [Sury, White et al. 2004].
The burning process results in organic pollutants like phenols, benzene, polycyclic
aromatic hydrocarbons, and heterocyclic compounds and inorganic compounds like
ammonia, cyanides, sulfates, and heavy metals [Humenick and Mattox 1978; Kapusta and
Stanczyk 2011]. The amount and nature of contaminants depends upon the type and rank
of coal ( hard coal, lignite etc.) [Kapusta and Stanczyk 2011]. These pollutants can move out
of the cavity either in aqueous phase or along with the gas leakage. However, this problem
can be rectified through careful site selection, appropriate operation control, proper shut
down process and effective environmental monitoring [Sury, Kirton et al. 2004]. In order to
avoid flow of contaminations from the cavity to the underground water table and to
minimize the loss of product gases, the pressure of burn cavity must be maintained below
hydrostatic pressure, to ensure a small and continuous influx of water into the cavity to aid
the burning process and to act as a barrier to contaminants outflow [Shafirovich and Varma
2009]. The Chinchilla project in Australia successfully demonstrated that UCG is an
efficient, economic and environment friendly process to exploit energy from coal. A
controlled shut down of the project in 2002-2003 revealed no ground water contamination,
no surface area disruption, no observable subsidence and no gas leakage [Scott and Steve
2006].
2.8 Uses of UCG
The product of UCG is a gas at high temperature and pressure. This is a combustible gas
and can be utilized for electricity generation in combined cycle gas turbine after minimal
processing. After further processing and refining it can be used for production of a wide
variety of gases, liquid fuels and chemicals [PWC. 2008]. Some uses of syngas are illustrated
in the Figure 2.9 [Courtney 2009].
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Virginia Tech | Chapter 2 - Literature Review
geology, type of oxidants used (air, oxygen, steam) and process parameters. Syngas can be
used directly as a fuel for boilers or electricity generators and as a feedstock for
manufacturing of chemicals. The composition of raw product gas is similar to that
produced by surface gasifiers, and cleaning technology for such gas compositions is already
a2v.8a.i2la bElel e[Cctrreiecd Pyo, Gwaerrn er et al. 2001].
Syngas, if not transformed, is mostly used as a fuel for electric generation. The clean gas can
be utilized in a gas-turbine based combined cycle power plant (IGCC). The integrated
gasification combined cycle power plants are a matured technology and are in use
worldwide. The IGCC power plants are being developed with the integration of carbon
capture and sequestration (CCS) to reduce the greenhouse gas emissions [Cormos, Starr et
al. 2010]. This means an IGCC plant based on UCG can provide all the benefits of IGCC
power generation with the added advantage of eliminating coal mining and surface
g2a.8si.3fi caStyionnt hpelatnict/ [SWuablskteirt,u Bteli nNdaetrumraaln G eat sa l(.S 2N0G0)1 ].
4
Another important use of syngas is to produce methane (CH ). Through the methanation
process, syngas can be converted into methane or synthetic natural gas, which can provide
an alternate domestic fuel supply utilizing the same infrastructure and end use equipment
[PWC. 2008; Oukaci 2009]. The process to convert coal/ coal gas to synthetic natural gas is
already developed and Dakota gasification company has been making SNG since 1980s
[PWC. 2008]. This SNG can utilize the available pipeline networks for transportation to
domestic and commercial consumers and can increase energy security especially in
countries where natural gas deposits are less than coal deposits.
Two major chemical reactions for production of SNG are Methanation and CO shift
reactions. These are exothermic reactions and can be used to produce steam [Oukaci 2009].
Methanation Reactions:
2 4 2
CO + 3H CH + H O + ΔH
2 2 4 2
CO + 4H CH + 2H O + ΔH
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2.8.7 Dimethyl Ether (DME)
DME, formed by the dehydration of methanol can act as an alternative fuel for
compression-ignition (diesel) engines. It has the handling characteristics of LPG,
x
completely sootless with 90% NO reduction and may compete favorably with conventional
f2u.8el.s8 [OGuaksa ctoi 2 L0i0q9u]i.d s
There are two commercially proven technologies for conversion of coal into transportation
fuels through gasification, Fischer-Tropsch process and Methanol to gasoline (MTG)
process. Both technologies follow similar route to convert coal into liquids, with changing
r2e.8ac.9ti onM ceotnhdaintiooln tso r Gesauslotlininge i n(MtoT vGar)i ous products [ExxonMobil 2009].
ExxonMobil has developed a technology to convert Methanol to high quality, ultra clean
compatible to refinery regular gasoline and commercially operated in New Plymouth, New
2Ze.8a.l1a0n dF i[sEcxhxeorn-MTroobpils 2c0h0 (9F]-. T )
The Fischer-Tropsch reaction converts a mixture of hydrogen and carbon monoxide,
derived from syngas in case of UCG, to liquid fuels. This process was discovered by German
scientists and used to make fuels during World War II. SASOL in South Africa has been
producing liquid fuels from coal for 30 years and many oil companies like Shell Oil,
Chevron (Texaco) and Exxon Mobil have been conducting research on this process [NETL
DOE. 2008]. UCG can provide feedstock for a F-T reactor and result in large amounts of
aviation and diesel fuel [PWC. 2008]. F-T is a mature industry and does not require any
specific distribution infrastructure, thus has market in the conventional fuel products
[Oukaci 2009]. Utilizing syngas as F-T feedstock opens its doors for advanced technologies
of converting syngas to transportation fuels, chemicals and domestic energy supplies.
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2.9 UCG Worldwide
2.9.1 USSR
The former Soviet Union started research on UCG as early as 1930s. For next four to five
decades, they reigned in this field. This technology matched their socialistic idea of
reducing the burden and working hours of labor force. It eliminated extensive and
painstaking underground mining and provided an effective source of energy. Lenin brought
the idea of coal gasification into public perception and Joseph Stalin started an extensive
research program [Klimenko 2009]. The first site was developed at Gorlovskaya in the
Donbass but was abandoned due to World War II. A second site was started in Lisichansk in
Donbass in 1935, reconstructed in 1949 [Olness and Gregg 1977].
In the 1940, the Russians started experimental work on brown coal at Tula in the Moscow
basin. In the1960s several thousand people were involved in UCG production but the
efforts declined rapidly in the 1970s due to the discovery of large natural gas resources
[Olness 1981]. They did not abandon the work entirely but continued at much reduced
rate. No new plants were set up during this period, however three stations were kept
running. The work at Angren, now in Uzbekistan, started in 1959 and the site is still in
operation [Brown 2012]. Similarly, Yuzhno-Abinsk, in Kuzbass Coalfield Siberia is also
currently operational. The Soviet work demonstrated that UCG could be applied in any
geological setting. They tried UCG on different coal ranks & types, thin to thick, shallow to
deep and flat to steeply dipping coal seams [Fair, Larson et al. 1976].
In earlier trials, Russians tried to burn the coal adjacent to the roadways and drifts. But it
resulted in the problems of air bypass and burning of much of the produced gas, hence they
started to use the boreholes which proved very successful in overcoming these problems
[Olness 1981]. Currently efforts are underway to bring UCG technology back in Russia and
some projects are in pipeline. VNIIPromgaz R&D center is currently trying to revive the
research in this field in Russia and implementing the technology abandoned by them to
fulfill the energy demands of the country [Kreinin 2006].
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Virginia Tech | Chapter 2 - Literature Review
2.9.2 The Europe
In several European countries, research on UCG continued between late 1940s and the
1980s, however major efforts for commercializing this technology started in the 1990s. The
main efforts of these trials focused on application of this technology on high rank, deep coal
seams, typical of Europe [Ag Mohamed, Batto et al. 2011]. Various trials took place in
B2.e9lg.3iu mU,K F rance, Spain and the UK, a summary of which is given below.
Sir William Ramsey planned the first test on UCG in County Durhum, UK in 1912. However,
concrete efforts for UCG started in the 1940s when experimental work started at Newman
Spinney, Derbyshire in 1949 and a second site opened at Bayton Worcestershire in the
following year [Thompson, Mann et al. 1976]. In 1990s the UK government supported the
European Union UCG research program in Spain and in 1999 started a new research and
development program [Couch 2009]. Currently Clean Coal and British Coal Gasification
(BCG) Energy Ltd are focused on developing deep coal reserves through the underground
gasification of coal and the generation of power from syngas and natural gas [Courtney
2008]. Thornton New Energy, a subsidiary of BCG Energy Ltd, in January 2009 received the
first UK license to carry out UCG and develop deep, previously un-minable coal reserves
under the Firth of Forth, Scotland and recently Thornton New Energy and Waste2Tricity
joined to convert coal into electricity combining new generation fuel cells with UCG [BCG
E2n.9e.r4g yF 2r0a1n0c]e.
In France, the work on UCG started in the late 1970s and continued until 1985 with
experiments at Bruay-en-Artois and La Haute Deule, in northern coalfield. In these trials
hydro-fracturing behavior, ignition characteristics and reverse combustion were analyzed
[Clean Coal Limited 2008]. In another trial at Echaux near ST Etienne, electro-linkage was
tried with limited success. Further development on UCG stopped because the experiments
resulted in the blockage of production wells by the formation of tars, oxidation products
and coal particles [Couch 2009]. No new developments have been reported in France since
1989.
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2.9.5 Belgium
In Belgium, an experiment to gasify semi-anthracite was conducted in 1948, at Bois-
la_Dame by Socogas. The coal seam was steeply inclined and the flame front advanced
sideways from an inclined drift rather than upwards from a horizontal one and probably
this approach contributed to poor results [Thompson, Mann et al. 1976].
In the 1980s, a UCG experiment was carried out at Thulin on a deep coal seam under the
Belgian-German joint research project. The coal was semi-anthracite, non-swelling,
multilayered with a cumulative thickness of about 6 m and depth of 860-870 m [Couch
2009]. A special drilling technique was used successfully to achieve linkage and CRIP was
used in one of the test [Scott and Steve 2006].
Belgium was also part of the Spanish trial, carried out in 1992-1999 under the European
C2o.9m.6m iSsspiaoinn a long with Spain and UK.
In El Tremedal Spain, UCG trial for deep underground coal seams was carried out as a joint
venture between Spain, UK and Belgium under the European Commission. The trial was
initiated at 1989 and finally abandoned in 1998. The trial evaluated the feasibility of UCG
for coal seams at depths between 500-700 m. In this trial the CRIP process was used and
linkage between wells was achieved by oilfield deviated directional drilling techniques
[Clean Coal Limited 2008]. The Spanish trial provided the basis for several other trials in
Europe and other countries, especially in the U.K. where the Department of Trade and
Industry Technology took initiatives for the development of UCG technology [Scott and
Steve 2006]. Another important aspect of the Spanish trial was implementation of the
lessons learned from the U.S. trial at Hoe Creek. The underground conditions were kept
close to hydrostatic pressure in order to avoid underground water influx and gas leakage
2th.9ro.7u ghO tthhee rs tErautrao [pCeraened Cyo, uGnartnrieers e t al. 2001].
In Europe, a large activity regarding UCG is going on. A project called “Hydrogen Oriented
Underground Coal Gasification for Europe” or HUGE is underway under the European
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Commission. Under the HUGE project, UCG trials will be carried out in an experimental
mine called “BARBARA” [HUGE 2010]. Interest in Poland, Hungary, Bulgaria and other
central European countries is growing and White Coal Energy & UCG Engineering Limited
have UCG projects in Hungary and Bulgaria [Courtney 2008]. UCG Engineering Limited has
contracts for UCG development projects in Eastern Europe and UK [UCG Engineering Ltd
2009]. Similarly Clean Coal Limited and UCG Partnership, in conjunction with other
partners and research groups, are supporting research for UCG and are actively engaged in
organizing conferences, seminars and workshops for promotion of UCG technology in
E2u.9r.o8p eU aSnAd other countries.
The idea of gasification either underground or above ground, using mined coal has very
long history in the U.S. and dates back to late 1800s when lamplighter used to make their
rounds in larger cities, lighting streetlights fueled by “town gas” [Lawrence Livermore
National Laboratory. 2007]. This “town gas” was crude form of coal gasification and
disappeared when large reserves of natural gas were discovered [Lawrence Livermore
National Laboratory. 2007]. In the late 1940s a small test was carried out at Gorgas,
Alabama but the main program started in 1973 [Klimenko 2009]. The main sponsor for this
program was the U.S. Department of Energy (DOE) with some industrial partnerships.
During 1973 to 1989, more than thirty tests were carried out with the focus on specific
engineering concerns such as improvement in the linkages between wells and the
permeability of coal [Couch 2009]. This $ 350 million program was a technical and
environmental success but it could not reach commercialization, partly because of dramatic
drop in oil and natural gas prices in the mid-1980s [Friedmann 2010]. The U.S. DOE
program was a well-planned series of trials aimed at exploring the performance and
behavior of UCG reactors. The tests investigated the effects of using different oxidants
(oxygen, air, steam) on syngas quality and yield, evaluated the proper shutdown process
and resulted in the development of CRIP process used in Centralia, Washington and rocky
Mountain 1 in Wyoming [Couch 2009]. Following is a summary of DOE trials given by Davis
[Davis and Beath 2006];
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•
Hanna, Wyoming, conducted by Laramie Energy Technology Center on sub bituminous
• coal in 1971-1981
Hoe Creek, Wyoming, conducted by Lawrence Livermore National Laboratory on sub
• bituminous coal in 1972-1982
Morgan Town, West Virginia, conducted by Morgantown Energy Technology Center on
• bituminous coal in 1976-1982
Centralia, Washington, conducted by Lawrence Livermore National Laboratory on sub
• bituminous coal in 1981-1985
Rocky Mountain 1, very close to Hanna site, conducted by an industrial consortium led
by Gas Research Institute on sub bituminous coal in 1986-1993
The trials financed and led by industrial consortiums include the following [Davis and
B• eath 2006]:
• Rockdale, Texas, led by Texas A&M University on lignite in 1978-1980
• Rocky Hill, Wyoming, led by Atlantic Richfield Co on sub bituminous coal in 1977-1982
Tennessee Colony, led by Texas Utilities on lignite in 1976-1980
Though not all these test were completely successful, as some of these resulted in the
contamination of potable aquifers especially in Hoe Creek and Carbon County Wyoming,
they provided very useful information about the engineering, technical and environmental
aspects of the UCG operations. One of the major contributing factors in the failure of UCG
pilots in Hoe Creek and Carbon County Wyoming, USA, was poor site selection [Clean Air
Task Force 2009]. These projects resulted in the contamination of potable ground water
and seriously hampered further developments of UCG in the U.S [Klimenko 2009] .
Although the U.S. DOE heavily funded the pilot researches in the 1970s and 80s, the
resulting environmental impacts blocked further funding into UCG. Another factor in the
abandonment of UCG trials was availability of cheap oil and natural gas in the 1980s
[Friedmann 2010].
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UCG research in the U.S. has resulted in the following contributions [Creedy, Garner et al.
2• 001]:
• Cavity growth modeling concepts
• Development of CRIP
• Guided drilling techniques to establish gasification channel
• UCG application on bituminous coal
• Environmental impact data of UCG
Importance of hydro-geological characterization
UCG efforts are reviving in the U.S. and several private companies and industrial
organizations are promoting research in this field. In 2006, DOE commissioned Lawrence
Livermore National Laboratory to evaluate the status of UCG technology. They prepared a
draft report “Best Practices” in 2006 and updated recently. They are also developing a
detailed model of contamination flow and transport specifically for UCG [The Working
G2.r9o.u9p oCnh iUnCaG 2007].
China has the largest UCG program currently underway with at least 16 trials carried out
since the late 1980s [Brown 2012]. Coal is the most important energy source in China and
there are nearly 300 abandoned mines and with some 30 Gt of coal resources [Couch
2009]. In China, the UCG is applied on abandoned mine galleries with access gained
through existing workings, in order to recover the remnant coal from mines. The China
University of Mining and Technology, Beijing has developed a long tunnel, large section,
two-stage method for UCG production capable of producing gas with the heating value in
3
the range of 12-14MJ/m . There is a UK-China technology transfer project in progress for
the development and commercialization of UCG [Creedy, Garner et al. 2001]. XinAo group is
working on a UCG project on sub bituminous coal in China. The group is planning to setup a
methanol plant based on UCG syngas feed [Wan and XinAo Group 2006].
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2.9.10 India
The Oil and Natural Gas Corporation Limited (ONGC) discovered large reserves of about
350 billion tonnes of coal at depths more than 600 m in Gujarat and Bengal and the
prevailing geological, hydro-geological and strata conditions favor the use of UCG for
exploitation of these reserves [The Working Group on UCG 2007]. The ONGC and the Gas
Authority of India Ltd. (Gail) have prepared feasibility of UCG on pilot basis and ONGC have
signed a Memorandum of Understanding with the Skochinsky Institute of Mining (SIM) of
Russia and Coal India Limited (CIL) for a UCG pilot study [Khadse, Qayyumi et al. 2007].
Indian government is in the process of making the policy to designate blocks for UCG in
order to facilitate the application of this technology [Khadse, Qayyumi et al. 2007]. The
Australian government is also supporting UCG development in India and helping in the
preparation of feasibility for UCG application in Andhra Pradesh. Cougar Energy Limited
has signed a UCG contract with Essar Exploration and Production Limited (EEPL) for UCG
d2e.9v.e1l1op Smouetnht iAnf rInicdaia [Walker 2008].
South Africa has been using coal to liquid technology since 1955, with most of this work
confined to surface gasification but is now promoting research to use low grade, unminable
coals for power generation and liquid fuel manufacturing through UCG [PWC. 2008]. Sasol,
South Africa is planning to invest in a pilot plant to investigate the commercial viability of
UCG in the CTL (coal to liquid) process [Hattingh 2008]. Another South African Company
Eskom, has developed a UCG pilot plant at Majuba Colliery to prove the ability of co-firing
with coal at Majuba power station [PWC. 2008]. This was the perfect location for UCG pilot
th
as the sub bituminous coal at mine cannot be mined conventionally [PWC. 2008]. On 28
October 2010, Eskom’s UCG demonstration plant delivered gas to Majuba Power Station,
and co-fired with coal to produce 3MW electricity [Eskom 2010]. This was the first
c2o.9m.1m2e Arcuiastl rparloiad u ction of electricity from UCG after the Soviet program.
In Australia, several projects are underway to establish the commercial viability of UCG.
The Chinchilla project is a milestone in the commercial development of UCG. As a joint
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venture of Linc Energy Ltd. and CS Energy Ltd., a pilot plant started in 1999 at Chinchilla,
350km west of Brisbane [Walker, Blinderman et al. 2001]. The burn started initially at
three process wells and expanded to eight operating wells [Walker, Blinderman et al.
2001]. The project ran from 1997 through 2003 with a controlled shutdown and restart
and was the largest in the West [Burton, Friedmann et al. 2006]. Ergo Exergy Technologies
Inc. provided the technology for the project and designed the UCG plant, under an
agreement with Linc Energy Ltd [Linc Energy]. This project demonstrated that UCG could
provide commercial quantities of industrial gas for power generation at an economically
viable price to ensure the competitiveness of coal based IGCC in the market [Walker,
Blinderman et al. 2001]. Linc Energy Ltd also planned to develop a small coal to liquid
(CLT) plant with the support from Syntroleum Corporation, a Fischer-Tropsch technology
provider [Davis and Beath 2006].
Currently Cougar Energy Ltd is working on a 400 MW project at Kingaroy Queensland, as
well as UCG projects at Wandoan Queensland and Latrobe Valley Victoria, in Australia
[2C.9o.u1g3a rP Eakneisrtgayn 2 010].
Pakistan has estimated coal reserves of about 185 billion tons, out of which 175 billion ton
reserves are located at Thar in the southern part of Sindh province. The coalbed thickness
in Thar ranges from 0.20-22.81 m with multiple seams reaching up to a maximum of 20
seams at some places with cumulative thickness of 36 m [Sindh Coal Authority 2005; Thar
Coal & Energy Board 2008] . The coal ranges from Lignite-B to sub-bituminous-A with high
moisture and low sulfur contents. Thar coalfield is part of Thar desert, covered by sand
dune that extends to an average depth of over 80 m and overlies alluvial deposits
containing sandstone, siltstone & claystone reaching up to 209 m [Sindh Coal Authority
2005]. Government of Pakistan is planning to apply UCG for exploiting Thar coal and has
taken several steps in this regard. A pilot plat to produce 50 MW electricity from Thar coal
through UCG is in progress and the first stage of UCG pilot is expected to be operational in
May 2011. Cougar Energy Ltd., Oracle Coal Fields Ltd. and China Petrochemical Corporation
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Virginia Tech | Chapter 2 - Literature Review
(Sinopec) are also conducting feasibility studies for UCG in Thar and hold leases in that
c2o.9a.l1fi4el Od t[hSerirvsa stava 2011].
Several other countries are interested in UCG development and either pilot projects or
feasibility studies are underway. Those examining UCG potential include, Bulgaria, Former
Czechoslovakia, Italy, Morocco, Oman, Poland, Romania, Thailand, Brazil and Turkey
[Creedy, Garner et al. 2001]. Clean Coal Limited and Linc Energy Limited have signed MoU
for UCG development in Indonesia and Vietnam respectively [Courtney 2008].
2.10 Challenges
UCG is a method that requires proper environmental management. Major environmental
problems with UCG include ground water pollution and subsidence. UCG may also create
hazards of atmospheric emissions and human impact such as noise, increased traffic and
dust [Creedy, Garner et al. 2001]. Sulfur and nitrogen also rise to the surface along with
2
product gas [Fergusson 2009]. Similarly, CO management may pose problems if the
process is not adequately planned. Although UCG is not a new technology, it is still a
developing technology and requires many research and development initiatives for
realization of its commercial potential. Research about environmental, economic and
technical aspects of UCG is needed to prove its maturity as a viable coal exploiting
technology in comparison to other competitive energy sources like CBM, natural gas,
surface gasification and conventional coal mining. These challenges provide opportunity
for the advancement and growth of this technology to provide its share in the energy mix.
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Virginia Tech | Chapter 3 – Operational Parameters of UCG
3.1 Ignition
Ignition of coal seam generally starts after the wells have been drilled and gasifier is ready
to operate. Several methods can be used for insitu ignition of coal seam but the most
popular method is utilizing electrical resisting heating. In this method, two electrical
lighters (generally 1 Kw) are lowered into the well along with a thermocouple that
measures the changes in the temperature [Thorsness, Hill et al. 1977]. After the lighters are
lowered in the well close to the coal seam, charcoal briquettes, wood chips or other
flammable material is placed at the bottom of the hole until the lighters are covered with
this material. At this moment, the controlled flow of air is directed at the lighters and
charcoal. The charcoal and wooden chips ignite in a few minutes and in turn ignite the coal
seam. This is demonstrated by changes in temperature indicated by the thermocouple.
Once the ignition has started, the gasification process starts and the injection of gases and
steam starts at the rate and pressure designed for the particular gasifier.
Another method used for initial ignition of coal seam is through explosive material. In this
method, propane, gasoline or lighting fluid is injected into the coal seam just prior to start
of ignition and a light explosive charge is exploded at the bottom of the hole. The flame
generated by the explosion starts the fire that is sustained through the injection of oxygen
or air. Wooden chips or other flammable material can also be dumped into the hole, if
required.
3.2 Wells Pattern
Gen erally, four types of wells are drilled to form the gasifiers. They are
i. Injection wells
ii. Production wells
iii. Instrumentation/monitoring wells
iv. Dewatering wells
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Virginia Tech | Chapter 3 – Operational Parameters
Injection wells are used for injecting the oxidizers i.e. air, gas or steam into the seam. They
have compressors or compressing stations installed around them and inject high-pressure
gases into the gasifiers. The production wells are used to extract the syngas form the
gasifier. The injection and production wells are sometimes collectively referred as process
wells [Blinderman and Jones 2002]. The instrumentation wells are scattered in the area as
per design of the cavity. They contain instruments for measuring pressures, temperature
and flow rates and monitoring the flows in or out of cavity. The dewatering wells are used
to dewater the seam either prior to gasification or during the gasification process. These
wells are also used to collect samples for measurement of water quality prior to, during
and after the gasification process.
3.3 Wells Spacing
The spacing and number of process wells in the gasifier depend upon coal thickness, coal
rank, geology of the area i.e. number of prominent fractures or folds in the area,
permeability of the coal seam, gasification rate and economics of the project. Number of
wells requirement for thicker seams is less as compared to number of wells required for
thinner seams, thus in thicker seams wells can be at greater spacing than those in thinner
seams. Most common well spacing lies between 18 m (60 ft.) to 30 m (~100 ft.). The
increase in well spacing decreases the drilling cost and price of gas. A sensitivity analysis of
well spacing and cost of gas however indicates that the cost benefits decrease rapidly after
the well spacing is more than 150 feet [Boysen 1978]. The reason is contribution of drilling
cost to the operating cost, which has reduced impact after this spacing. Monitoring wells
are generally placed around 15 m and 30 m radius [Snoeberger 1977].
3.4 Wells Structure
The process wells are generally steel cased and cemented to withstand the higher
temperatures and pressures. These wells can be completed in the upper portion of seam
and perforated at the lower or bottom part of the seam [Fischer, King et al. 1977]. The
diameter of process wells is in the range of 14 to 20 inches [Boysen 1978]. The
instrumentation wells are generally not cemented or cased and are small diameter holes.
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Virginia Tech | Chapter 3 – Operational Parameters
They carry thermocouples and tubes for measuring temperatures, water levels and gas
sampling. The dewatering wells are steel cased to the top of coal seam and are perforated
within the coal seam and have steel pump sections that carry water pumps to collect water
from sumps [Thorsness, Hill et al. 1977].
3.5 Instrumentation
The instruments are used to measure temperatures, air and gas pressures, water level, flow
rates, and for gas sampling. Thermocouples are used to measure the temperature
variations within the coal seam and in the surrounding strata. A single instrument hole can
carry several thermocouples at different depths. They are generally placed within the seam
at varying depths (top, center and bottom of seam ) and at 5, 10, 15 and 20 feet in the over
and under burden to have complete temperature profile of the gasifier and surround rocks
[Thorsness, Hill et al. 1977]. The air and gas flow meters, transducers and gas
chromatographs measure the flow rate, pressure and composition of gas. Submersible
pumps are used in dewatering wells.
3.6 Well Linkages
The most important step during the creation of insitu gasifier is the development of linkage
between the injection and production well. This can be achieved through any of the
follo wing method.
i. Explosive fracturing
ii. Reverse combustion linkage
iii. Hydraulic fracturing
iv. In seam channels through drilling
v. CRIP (Controlled Retraction Injection Point)
In the explosive fracturing technique, high explosive is detonated at the bottom of the hole,
either in one of the process wells or at a specially drilled well for this purpose. The idea
behind this technique is that explosive fracturing will enhance the permeability of the coal,
hence the linkage between the wells. An experimental blast consisting of two high
explosive spherical shots at Hoe Creek Wyoming revealed that this technique creates three
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Virginia Tech | Chapter 3 – Operational Parameters
regions after explosions [Thorsness, Hill et al. 1977]. The region where there is no change
in the permeability and the permeability value remains as it was before explosions. This
region is at a radial distance of about 50 feet from explosions depending upon type of
explosive and geology of the area. A very high permeability area within 10 feet of explosive
charges and an intermediate area where permeability is significantly enhanced after
explosions. However, this technique is not very successful as in the areas very close to
explosion, the permeability actually decreased due to plugging of fractures by fine coal
particles resulting from the blast [Thorsness, Hill et al. 1977].
Reverses combustion linkage is somewhat more successful method than explosive
fracturing and has been tried in several experimental trials and pilots in the Europe, the
U.S. and other places [Zamzow 2010]. In the reverse combustion linking, the fire is started
at one well, well A in Figure 3.1. The air is initially injected in the ignition well (well A) to
sustain the fire. Once the combustion of coal is started, the air injection is switched and
injected to the other well, well B in Figure 3.1. Now the air current moves towards the
ignition front through the coal seam and the combustion moves from ignition well (well A)
to the injection well (well B). The combustion zone is moving against the current of air
towards oxygen source (well B) and because of this movement against the air current, the
process is called reverse combustion. Because of this movement of combustion zone
towards injection well, a highly permeable localized zone is created in the seam. After the
combustion zone reaches the injection well, the reverse combustion is completed [Fischer,
King et al. 1977]. At this point high volume, low-pressure air is injected into the injection
well (well B) and combustion zone now moves towards ignition well (well A). At this point
air current and combustion zone move in the seam direction and this is called forward
combustion. The position of reverse combustion link at the bottom of seam is very
important, as in the forward combustion this undercut zone will result in falling of fresh
coal in the reaction zone, yielding high efficiency and producing a packed bed [Fischer, King
et al. 1977].
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Virginia Tech | Chapter 3 – Operational Parameters
Figure 3.1: Reverse combustion linkage
Hydraulic fracturing is another technique used to improve the linkage between wells. In
this process, the low-pressure water is first used to washout the borehole. The pressure is
then increased gradually until the cracks are generated in the seam. After that, hydraulic
fluids are pumped into the cracks. Finally the area is flushed with water and then with air
to remove any plugging [Olness 1981]. The control in hydraulic fracturing is very
important, as the process can increase the potential of gas losses greatly due to irregularity
of fractures development.
In-seam channels are one of the better options to improve well linkages. In this method, an
in-seam cannel is created by drilling a horizontal borehole in the lower part of coal seam.
Directional drilling reduces the surface disruption and increases the amount of coal that
can be accessed by a pair of wells [Ahner 2008].
The CRIP technology is a modern technique, giving excellent control in the cavity formation
and linking the wells and is based on advancements in directional drilling. CRIP stands for
Controlled Retraction of Injection point. In this technique, the injection point is retracted
occasionally to expose the fresh coal to the injecting gases. In this method, the link between
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Virginia Tech | Chapter 3 – Operational Parameters
the wells is established through a directionally drilled horizontal borehole between the
injection and projection well. After a cavity has been formed in the coal at a specific
location, the injection point is retracted either mechanically or by burning the portion of
the injection tube, thus exposing the fresh coal to the stream of gases and developing a new
cavity. 10 to 20 such movements can occur during the lifetime of a set of wells [Clean Coal
Limited 2007].
3.7 Operating Pressures
The operating pressure required for gasification depends upon the type of coal, the
permeability, depth of seam, type of oxidant used, amount of water and moisture in the coal
and strata, and most importantly the hydrostatic pressure around the burn cavity. The
experiments in the U.S. dictate that the operating pressure should never exceed the
hydrostatic pressure of the cavity, otherwise the hazardous waste and pollutants can
migrate out of the cavity and can result in the contamination of surrounding area and water
resources [Ahern and Frazier 1982; Younger, González et al. 2010]. The air pressure also
depends upon the method used for linking the wells. For reverse combustion and hydraulic
fracturing high pressure is required to move the ignition front whereas low pressure is
required for CRIP and in-seam borehole. The gas pressures at the beginning of ignition are
high to promote combustion of coal, but once combustion is achieved, injection pressure is
reduced and flow rate is increased. Typical injection pressures range between 60 to 120 psi
and the flow rates vary from around 200 scfm in the beginning to around 4500 scfm at
various stages of gasification [Fischer, King et al. 1977].
3.8 Cavity Development
After ignition, the cavity starts to develop along the line of linkage between the wells. In
case of reverse combustion linkage, the cavity develops in the form of a cylinder with the
final radius almost equal to the distance between the process wells [Fischer, King et al.
1977]. The initial development of cavity is in the horizontal direction but later the cavity
developments occurs in a vertical direction and burning g takes place near the top of seam
[Roehl, Brown et al. 1977]. Thus, the initial linking path should be close to the bottom of the
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Virginia Tech | Chapter 3 – Operational Parameters
seam during reverse combustion to get maximum recovery in forward combustion. For
directional drilling and CRIP, the cavity develops around the borehole and expands
laterally.
3.9 Water Control
Water plays a crucial role in the gasification process. The presence of water affects the
nature and quality of the product gas [Walters and Nuttall 1977]. The presence of water is
necessary for sustaining the reaction and water inflow of about 1 gpm is considered good
for gasification [Snoeberger 1977]. The water inflow depends upon the presence of aquifer
in the area and permeability of the seam. In high permeability area, the water influx can be
higher than the desired levels and can decrease the heating value of the gas, as vaporization
of the water consumes part of thermal energy. Excessive water influx can be controlled by
using the dewatering wells, controlling the cavity or gas pressure, gasifying the seam in an
up dip direction and two-stage pyrolysis gasification[Gunn 1977; Vanderborgh, Wewerka
2
et al. 1977] . In the two-stage gasification pyrolysis process, CO is recovered from the
product gas, heated and injected into the coal seam prior to gasification in that portion.
2
This heated CO dries the coal seam, thus resulting in less water contents and increased
porosity of coal because of the water removal from pores [Vanderborgh, Wewerka et al.
1977] .
3.10 Gas Cleaning
The level of cleaning required for product gas depends upon the amount of sulfur
compounds, moisture contents, particulate matter and gas composition. Sulfur and
nitrogen report to the surface with the product gas but sulfur can be removed using sour
gas cleaning technologies that are already available and quite common [Fergusson 2009].
2
The CO removal from syngas stream is quite easy and economical and technologies are
2
currently available to separate CO from syngas either onsite or at power plants in a cost-
effective manner [Clean Air Task Force 2009]. The gas stream exiting the production well is
at elevated temperatures and gas cleaning is accomplished more easily if the gas is at lower
temperatures [Vanderborgh, Wewerka et al. 1977]. This is achieved through the
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Virginia Tech | Chapter 3 – Operational Parameters
introduction of a heat exchanger in the system before transporting gas to the gas cleaning
facility. If the gasification is taking place at lower depths and the pressure of syngas is less
than the operating pressures required for turbine, a compressor can be inducted into the
process line to pressurize the gas at desired levels [Blinderman and Jones 2002]. However,
for gasification of deep-seated coal seams, the compressor is not required as product gas is
already at elevated pressures. Cleaning of product gas from UCG is easier and relatively less
expensive than that from the surface gasifiers because the more viscous coal tars from
surface gasifiers plug the valves and other equipment, whereas tars form UCG have low
viscosity similar to that of oils [Gunn 1977] and are easy to remove for gas cleaning.
3.11 Economics of the Process
The economics of UCG largely depend upon the type of coal, depth and thickness of coal
seams, number and diameter of wells, well completions type and extent, the linking
technique used, type and amount of oxidant used, gas pressure, quality, heating value,
leakage, subsidence, water influx rate, type and size of compressors used, gas cleanup
requirements, type and number of instruments and conversion and recovery efficiencies.
Seam thickness and well spacing have more pronounced impact on gas prices as compared
to other variables [Boysen 1978]. The cost of syngas is comparable to very low cost thermal
coal and the cost of electricity generation from syngas is very competitive to coal fired
power stations. The capital cost of UCG_IGCC is almost half of that for coal-IGCC plant as no
surface gasifier is required and UCG-IGCC can produce electricity at a price as low as
$10.00/MWh [Blinderman and Jones 2002]. A detailed economic analysis by Indiana
University for UCG application to Indiana Coal indicates that production cost for UCG based
on air as oxidant is $8.04 per MMBtu and for oxygen oxidant, this cost drops to $
4.8/MMBtu. The cost of electricity production at combined cycle plants is $ 0.0863/KWh
for air-fired UCG and $.0643/KWh for oxygen-fired UCG [Ag Mohamed, Batto et al. 2011].
The carbon capture and sequestration adds 1.7 cents per kilowatt hour and makes UCG
more economical than competing technologies in the case when carbon tax or cap and
trade is imposed [Ag Mohamed, Batto et al. 2011].
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Virginia Tech | Chapter 4 – Development of Site Selection Criteria for UCG
4.1 Introduction
UCG site selection depends upon a number of parameters related to coal, strata, water
conditions and surface facilities. Most important parameters are coal rank, seam thickness
and depth, type of overlying and underlying strata, permeability and porosity of coal and
strata, geological features such as faults, joints, beddings and fracture networks, location of
potable aquifers & their composition and closeness to surface infrastructure.
This chapter presents a scheme for both resource assessment and selection of potential
UCG sites in central Appalachia, based on a detailed set of site selection parameters. A case
problem approach is used to show how these different parameters can be applied to the
available data to identify potential UCG sites. In addition, the concepts developed in this
research can be used to delineate requirements and criteria for UCG site selection.
The Central Appalachian Basin encompasses approximately 10,000 square miles in
southwestern Virginia, southern West Virginia and eastern Kentucky. It is the most mature
and extensively mined coal-producing region in the United States. However, difficult
mining conditions, stringent environmental regulations, strong competition from other coal
producing regions and other energy sources and depletion of most accessible and low cost
reserves, have contributed to declining coal production in central Appalachia. As a result, it
is important to examine alternatives to traditional mining for the region, including
opportunities offered by UCG. This technology has the potential to harness energy from
deep, thin, low-grade and unminable coal seams, thus greatly enhancing the availability of
exploitable coal reserves.
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Virginia Tech | Chapter 4 - Site Selection Criteria
4.2 Background
UCG is a process involving burning of coal insitu and bringing up the product in the form of
heated gases commonly known as syngas. The renewed interest in commercial
develop•m ent of UCG technology can be attributed to the following factors.
•
stringent environment and safety regulations specific to coal mining
•
instable and politically motivated increases in petroleum prices
•
increasing negative public perception about coal mining
•
rapidly increasing mining costs
•
decrease in economically minable resources
•
environmental and economic potential of UCG
•
ability to recover energy from previously mined coal seams
applicability of UCG on low heating value, deep and relatively thin coal seams,
•
considered economically unminable through conventional mining methods
lower capital cost due to elimination of coal mining, transportation and ash
•
management facilities
reduced surface footprint
In order to realize the full economic and environmental potential of UCG, proper site
selection is a key factor [Sury, Kirton et al. 2004]. One of the major contributing factors in
the failure of UCG pilots in Hoe Creek and Carbon County Wyoming, USA, was poor site
selection [Clean Air Task Force 2009]. These projects resulted in the contamination of
potable ground water and seriously hampered further developments of UCG in the U.S
[Burton, Friedmann et al. 2006]. Although the U.S. DOE heavily funded the pilot researches
in the 1970’s and 80’s, the resulting environmental impacts (resulting due to poor site
selection and process control) blocked further funding into UCG. This emphasizes the
importance of a well thought-out resource assessment plan for development of UCG sites.
4.3 Sıte Selectıon/Assessment Crıterıa
The literature on UCG indicates that pilot projects and trials have been conducted for
several coal ranks, depths, seam conditions, geological and hydrological settings with
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Virginia Tech | Chapter 4 - Site Selection Criteria
different success levels. Over the last 100 years, experiments have been carried out on
coals varying from lignite to anthracite, shallow depths to deep lying and horizontal to
steeply dipping coal seams. From these studies, the following parameters are considered
the mos•t important for the selection of a proper UCG site:
•
Coal Rank/Type
•
Seam thickness
•
Seam depth
•
Seam inclination
•
Seam structure
•
Permeability and porosity
•
Moisture contents
Hydrogeology of the area and ground water issues
•
Other relevant parameters include:
•
Quantity of resources
•
Available infrastructure
•
Presence of coalbed methane
Land use restrictions and other regulatory requirements
In this chapter, the most relevant parameters for site selection have been reviewed and
applied for selection of sites in the Appalachian region of the U.S. to analyze the potential
f4o.3r .a1p pCliocaatli oRna nofk U CG.
All types of coal ranging from lignite to anthracite can be gasified [Burton, Friedmann et al.
2006], as demonstrated successfully in surface gasifiers and tested for UCG. However, some
type of coals are easier to gasify than others. For example swelling and coking coals can
block the linkages between injection and production wells upon heating and can result in
decreased permeability and porosity of coal [Couch 2009]. Testing on anthracite and semi-
anthracite was conducted in Former Soviet Union and Bois-la-Dame Belgium [Thompson,
Mann et al. 1976; Burton, Friedmann et al. 2006]; however, the results were not very
encouraging.
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Virginia Tech | Chapter 4 - Site Selection Criteria
Low rank coals especially lignite and sub-bituminous are preferable for UCG as the matrix
tends to shrink on heating, resulting in enhanced permeability and improved linkages
between injection and production wells [Bialecka 2009]. Low rank coals are preferred for
gasification because they are at relatively shallower depth and easy to ignite and they do
not form hard coke residue and are softer which makes it easy to develop a gasifier
[Thompson, Mann et al. 1976]. The other reason for giving less preference to anthracite for
UCG is its premium value. If anthracite for some reason is not minable through
conventional mining methods, then it may be considered for gasification [Thompson, Mann
et al. 1976]. Thus, the literature advocates that sub bituminous, non-coking coals are most
suitable for gasification.
A study of coal rank trends by Hower and Rimmer shows that coal ranks generally increase
towards the southeast in the Central Appalachian Basin in southern West Virginia, western
Virginia, and eastern Kentucky but the ranks decrease towards the Alleghany front in
Virginia and West Virginia [Hower and Rimmer 1991]. The study utilized volatile matter as
an important parameter in ranking coals. The Pocahontas No.3 coal in southern West
Virginia, is ranked as low volatile bituminous in McDowell and Wyoming counties. In
central Buchanan County Virginia, the Pocahontas No.3 is ranked as low volatile
bituminous, with volatile matter as low as 17% [Hower and Rimmer 1991]. However, the
Burtons Ford coalbed to the south in Russell and Scott Counties, Virginia, has a high volatile
A rank and coals above the Pocahontas No. 3 exhibit parallel trends, though at lower rank
levels [Hower and Rimmer 1991].
Conrad et al. prepared a coal rank map for the Central Appalachian Basin based on
percentage volatile matter which indicates the coal rank increases from high-volatile A
bituminous (lower rank) in the northwest and west to semi-anthracite (high rank) in
McDowell and Wyoming Counties, West Virginia [Conrad, Miller et al. 2006]. Overall,
majority of coal in Central Appalachia ranks between sub-bituminous to bituminous with
some smaller areas of semi-anthracite.
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Virginia Tech | Chapter 4 - Site Selection Criteria
4.3.2 Seam Thickness
Coal seam thickness is another important site selection parameter. Initially there was a
hope that UCG may effectively be applied to very thin coal seams that were not recoverable
through conventional mining methods. However, it was found that UCG works best for
thicker coal seams, especially seams more than one meter thick [Lauder and Smith 2012].
Although Russians tried to gasify coal seams less than one meter thick, the heating value of
gas was very low for these seams [Burton, Friedmann et al. 2006]. The decrease in the
heating value in thin seams may be attributed to the more pronounced heat losses to the
strata and difficulty in controlling the shape of gasifier passages, which may become wide
and allow the air to bypass reaction fronts and burn the forming syngas [Thompson, Mann
et al. 1976].
Although Ergo Exergy states that coal seams as thin as 0.5 m may be gasified through UCG
[Shafirovich and Varma 2009], a coal seam thickness of more than 2.0 m is preferable.
Thicker coal seams (> 2m) require fewer boreholes to gasify more amount of coal,
providing both economic and environmental advantages [Sury, Kirton et al. 2004]. In
thinner seams the inert floors and roofs act as heat sinks whereas in thick seams the
channel is in the same cross sectional area and the entire surface area is reactive, thus
there is no heat sink and all heat flows react on coal. Moreover the thick coal seam prolongs
the working life of gasifier [Thompson, Mann et al. 1976]. The lower limit for the seam
thickness should be more than 2 m, with the ideal thickness around 5 m to 10 m. This ideal
thickness provides significant quantities of coal to be gasified without involving
unacceptable ground movements and subsidence [Couch 2009]. The upper limit for coal
thickness is usually dependent upon geomechanical properties of the surrounding strata
and potential of cavity collapse. Russian projects indicate no maximum limit to seam
thickness [Burton, Friedmann et al. 2006], but with a seam thicker than 20 m, there is a
greater risk of subsidence and cavity collapse [Couch 2009]. Thus, most suitable seams may
lie between a thickness of 2m and 20 m.
The Central Appalachian Basin consists of coal seams that are comprised of widely
distributed multiple coalbeds. These coal seams from oldest to youngest are; the
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Virginia Tech | Chapter 4 - Site Selection Criteria
Pocahontas No. 3, Pocahontas No. 4, Fire Creek/Lower Horsepen, Beckley/War Creek,
Sewell/Lower Seaboard, and Iager/Jawbone [EPA. 2004]. The Pocahontas coal seams
include the Squire Jim and Nos. 1 to 7 with Nos. 3 and 4 the thickest and most areally
extensive [EPA. 2004].
Appalachian coal seams are geologically and technically challenging because numerous thin
seams (less than 1 m) are distributed throughout thick stratigraphic sections [SECARB.
2011]. Marshall Miller and Associates Inc. (MM&A) has developed various coal isopach
maps for different fields of the Central Appalachian Basin targeting the Lee and Pocahontas
formations. A net coal isopach for the Frying Pan field, Virginia representing composite coal
thickness of the Lee and Pocahontas formations (Greasy Creek to Pocahontas No. 1 coal
seams) indicates a net coal thickness of 6.9 m (22.5 feet) [SECARB. 2011]. The average net
coal thickness for the Sourwood field, Virginia in the Lee and Pocahontas formations is
about 8.0 m (26.4 feet), whereas the average net coal thickness in the Lick Creek field,
Virginia for the same formations is 5.7 m (18.6 feet) and for the South Oakwood filed,
Virginia it is 7.4 m (24.4 feet) [SECARB. 2011]. The Beckley seam in the Loup Creek field,
West Virginia has an average thickness of 1.6 m (5.3 feet); however, it ranges from 0.5 to
2.9 m at places [SECARB. 2011]. Figure 4.1, Figure 4.2, Figure 4.3 and Figure 4.4 show the
net coal isopach for different fields in the Basin, adapted after SECARB 2011.
Figure 4.1: Net coal isopach for Frying Pan field, VA
53 |
Virginia Tech | Chapter 4 - Site Selection Criteria
Figure 4.5 shows a net coal thickness isopach adopted from Ripepi indicating a net coal
thickness varying from less than one meter to more than 10 m [Ripepi 2009]. However, this
is the composite thickness of various coal seams in the region especially in the Lee and
Pocahontas formation. The individual seams in this region are generally thin (< 1 m) with
some thicker seams ranging to more than 2 m in thickness.
Figure 4.5: Net coal thickness isopach for Central Appalachia
Any typical cross section of the area indicating thickness of different coal members will
show that there are multiple coal seams in the region with varying depths and thickness.
The composite thickness of these seams reaches up to more than 10 m (~ 30 ft.) in some
areas, however, individual seams are generally thin but some are thicker than one meter
and can be potential UCG sites. A study is needed to see the applicability of UCG on multiple
seams in the region especially the ones that are relatively closer to each other. Figure 4.6
suggests a schematic of this idea.
55 |
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