&EPA
United States
Environmental Protection
Agency
Industrial Environmental Research EPA-600/7-79 1 24
Laboratory ju|y 1979
Cincinnati OH 45268
Research and Development
Dewatering Active
Underground
Coal Mines
Technical Aspects and
Cost-Effectiveness
Interagency
Energy/Environment
R&D Program
Report
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RESEARCH REPORTING SERIES
Research reports of the Office of Research and Development, U.S. Environmental
Protection Agency, have been grouped into nine series. These nine broad cate-
gories were established to facilitate further development and application of en-
vironmental technology. Elimination of traditional grouping was.consciously
planned to foster technology transfer and a maximum interface in related fields.
The nine series are:
1. Environmental Health Effects Research
2. Environmental Protection Technology
3. Ecological Research
4. Environmental Monitoring
5. Socioeconomic Environmental Studies
6. Scientific and Technical Assessment Reports (STAR)
7. Interagency Energy-Environment Research and Development
8. "Special" Reports
9. Miscellaneous Reports
This report has been assigned to the INTERAGENCY ENERGY-ENVIRONMENT
RESEARCH AND DEVELOPMENT series. Reports in this series result from the
effort funded under the 17-agency Federal Energy/Environment Research and
Development Program. These studies relate to EPA's mission to protect the public
health and welfare from adverse effects of pollutants associated with energy sys-
tems. The goal of the Program is to assure the rapid development of domestic
energy supplies in an environmentally-compatible manner by providing the nec-
essary environmental data and control technology. Investigations include analy-
ses of the transport of energy-related pollutants and their health and ecological
effects; assessments of, and development of, control technologies for energy
systems; and integrated assessments of a wide range of energy-related environ-
mental issues.
This document is available to the public through the National Technical Informa-
tion Service, Springfield, Virginia 22161.
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EPA-600/7-79-124
July 1979
DEWATERING ACTIVE UNDERGROUND COAL MINES:
TECHNICAL ASPECTS AND COST-EFFECTIVENESS
by
W. A. Wahler & Associates
Palo Alto, California 94303
Contract No. 68-03-2366
Project Officer
S. Jackson Hubbard
Extraction Technology Branch
Industrial Environmental Research Laboratory
Cincinnati, Ohio 45268
INDUSTRIAL ENVIRONMENTAL RESEARCH LABORATORY
OFFICE OF RESEARCH AND DEVELOPMENT
U.S. ENVIRONMENTAL PROTECTION AGENCY
CINCINNATI, OHIO 45268
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DISCLAIMER
This report has been reviewed by the Industrial Environmental Research
Laboratory, Cincinnati, U. S. Environmental Protection Agency, and approved
for publication. Approval does not signify that the contents necessarily
reflect the views and policies of the U. S. Environmental Protection Agency,
nor does mention of trade names or commercial products constitute endorsement
or recommendation for use.
ii
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FOREWORD
When energy and material resources are extracted, processed, converted,
and used, the related pollutional impacts on our environment and even on our
health often require that new and increasingly more efficient pollution con-
trol methods be used. The Industrial Environmental Research Laboratory -
Cincinnati (lERL-Ci) assists in developing and demonstrating new and improved
methodologies that will meet these needs both efficiently and economically.
One possible means of controlling acid mine drainage is to intercept
groundwater before it actually enters the active coal mine. This report
presents the technical results of a pilot-scale dewatering operation and
assesses the cost-effectiveness of the technique. The study should provide
the coal mining industry with useful concepts and general guidance in deter-
mining the technical and economical feasibility of thus dewatering an active
coal mine. The study also suggests measures that may be cost-effective in
controlling mine water quality with typical mine water removal systems. The
practical geohydrologic experience with wells gained during this study should
be useful to others involved with mine dewatering research and also points
out areas where additional work is needed.
Further information on this subject may be obtained from the Extraction
Technology Branch.
David G. Stephan
Director
Industrial Environmental Research Laboratory
Cincinnati
iii
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ABSTRACT
This study evaluated the cost-effectiveness of dewatering an active under-
ground coal mine as an alternative or supplement to treating acid mine drainage.
Through a combination of research, personal contacts, and site visits to se-
lected wet mines in the northern part of the eastern bituminous coal mining
region, the Barnes and Tucker Company's Lancashire No. 20 mine, located in
western Pennsylvania, was selected as a suitable site for a pilot-scale de-
watering operation. A dewatering program was formulated and base-line data
collection was performed in conjunction with exploration of hydrogeologic
conditions and well construction. Dewatering was then performed for a total
of 25 days to provide the basis for the technical and cost analysis.
The results of well testing and pilot dewatering indicated that ground-
water flow into the mine is controlled by fracture zones, a condition that
appeared to be prevalent in the wet mines inspected during this study. The
hydraulic characteristics of this type of flow system are difficult to define,
and a study is currently being performed by D'Appolonia Consulting Engineers,
Inc. to locate and define the quantities of water that enter underground mines.
However, the dewatering operation demonstrated that mine inflows could be in-
tercepted (well-effectiveness was projected to range from 50 to 80 percent)
and could reduce the amount of water degraded in the mine. Though individually
pumped wells constructed from the surface proved not to be cost-effective at
the study mine, this system could be cost-effective (without considering other
benefits such as roof stability), if the acidity of the mine drainage were
higher (ranging from 500 to 1500 mg/£ as CaC03), and if the intercepted ground-
water could be discharged without treatment. A more economical system (which
would be cost-effective at the study mine) appears to be the use of angled
drain holes drilled from the mine openings; costs for this system are not as
sensitive to drainage quality or mine depth. Monitoring of water quality as
it was transferred through the mine indicated that degradation is a gradual
process; therefore closed discharge systems or more direct transfer to the
surface might also prevent substantial water degradation.
This report was submitted in fulfillment of Contract No. 68-03-2366 by
W. A. Wahler & Associates, of Palo Alto, California, under the sponsorship
of the U.S. Environmental Protection Agency. This report covers the period
December 8, 1975 to June 30, 1978, and work was completed as of August 30, 1978.
iv
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CONTENTS
Foreword iii
Abstract iv
Figures vii
Tables ix
Symbols and Abbreviations x
Conversion Table xi
Acknowledgments xiii
Section
1 Introduction 1
2 Conclusions , 3
3 Recommendations 6
4 Background 7
General hydrogeologic conditions . . 7
Hydrologic effects of underground mining 8
Problems associated with mine water 10
Current mine water control 12
5 Site Selection 13
Selection criteria 13
Selection process 15
Development of a pilot dewatering plan 19
6 Site Conditions 20
Description 20
Investigative acitivies 28
Hydrogeology 37
Baseline data 46
7 Pilot Dewatering Program 58
Method of approach 58
Development 58
Operation 58
Monitoring 59
Results 62
8 Mine Dewatering Systems and Alternatives to Dewatering. . 77
Wells constructed from the surface 77
Wells constructed underground 78
Design approach 79
Alternatives to mine dewatering 80
9 Cost-Effectiveness of Dewatering 82
Introduction 82
Cost of mine water drainage and treatment -
Study mine 85
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CONTENTS (continued)
9 Cost-Effectiveness of Dewatering
Industry costs 99
Costs of mine dewatering using pumped wells .... 106
Cost-effectiveness comparisons 107
References 117
Bibliography 119
vi
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FIGURES
Number
1 Schematic asymmetry of drawdowns along a fracture zone 9
2 Location map: Lancashire No. 20 mine and pilot well
dewatering site 21
3 Water transport system 24
4 Average daily flows to treatment plant 25
5 Treatment plant - schematic layout 26
6 Main G study area 27
7 Mine water flow path and monitoring system - Main G study
area to treatment plant 29
8 Sampling box at G-ll sump (schematic) 31
9 Revised mine water flow path and monitoring system - Main G
study area to treatment plant . . . 33
10 Location map: Dewatering and observation wells 34
11 Areal geology 39
12 Site stratigraphic column 40
13 Site cross-section 41
14 Well locations and water contours 9/15/77 45
15 Average mine inflow - Main G study area 9/26/76 to 6/5/77 .... 47
16 Drilling log: Dewatering well P-l 51
17 Drilling log: Dewatering well P-2 52
18 Drilling log: Dewatering well P-3 53
19 Drilling log: Dewatering well P-4 54
20 Drilling log: Observation well OB-1 . 55
21 Drilling log: Observation well OB-2 56
22 Drilling log: Observation well OB-3 57
23 Location map: Surface water sampling stations 61
24 Average daily mine inflow - July 1977 dewatering 63
25 Average daily mine inflow - September 1977 dewatering 64
26 Average daily pumping rates - September 1977 dewatering 65
27 Combined daily pumpage rate - September 1977 dewatering 65
28 Water level elevations in dewatering wells -
September 1977 dewatering period 66
29 Drawdown: Observation well OB-1 - September 1977 dewatering. . . 67
30 Drawdown: Observation well OB-2 - September 1977 dewatering. . . 67
31 Drawdown: Observation well OB-3 - September 1977 dewatering. . . 68
32 Drawdown: Observation well P-2 - September 1977 dewatering ... 68
33 Operating costs for hydrated lime treatment facilities 83
34 Summary of cost analysis. , 98
35 Capital cost of treatment plant vs. plant capacity - hydrated
lime treatment with sludge disposal ..... 103
vii
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FIGURES (continued)
Number Page
36 Total operating cost (including capital costs) -
hydrated lime treatment 104
37 Operating and capital costs per well vs. well depth 109
38 Annual operating costs vs. number of wells and well depth. . . . 110
39 Capital investment vs. number of wells and well depth Ill
viii
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TABLES
Number Page
1 Water quality - Lancashire No. 24D mine 16
2 Water quality - Rushton mine 17
3 Water quality - Lancashire No. 20 mine 18
4 Baseline water quality data 50
5 Summary results - water quality 73
6 Pilot program: Water quality data - well waters 74
7 Pilot program: Water quality data - receiving stream 75
8 Pilot program: Water quality data - mine waters 76
9 Accounting methods in the coal industry 84
10 Summary of drainage annual operating expenses 87
11 Summary of water treatment annual operating expenses 88
12 Work orders - Water treatment and drainage - Fiscal 1976 .... 89
13 Summary of items capitalized as depreciable assets -
drainage and water treatment 90
14 Short summary of capital investment 91
15 Short summary of annual operating expenses 92
16 Annual mine cost of electricity for underground drainage .... 94
17 Mine cost summary 95
18 Underground drainage annual operating expenses -
cost contributors 95
19 Water treatment annual operating expenses - cost
contributors (1976 fiscal year) 97
20 Summary of capital costs - conventional lime
neutralization process 100
21 Summary of operating costs - conventional lime
neutralization process 101
22 Comparison of cost contributors: Study Treatment
facility vs. operation Yellowboy results 102
23 Capital costs for three wells 152 meters (500 ft) deep 107
24 Annual operating costs for three wells 152 meters
(500 ft) deep 108
25 Pilot dewatering costs vs. mine drainage and treatment costs . . 112
26 Active mine area dewatering costs vs. mine drainage and
treatment costs 113
27 Capital costs for inclined drainage holes . 116
ix
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SYMBOLS AND ABBREVIATIONS
SI Metric
Symbol Unit :
m meter
nm nanometer
cm centimeter
kilometer
square meter
hectare
cubic meter
second
day
meters per second
cubic meters per second
cubic meters per meter per day
liter
liters per second
gram
milligram
kilogram
milligrams per liter
t tonne
W watt
kW kilowatt
yS/cm microsiemens per centimeter
Customary U.S. Measure
ha
m3
s
d
m/s
m3/s
m-Vm-d
I
l/s
g
mg
kg
Abbrev. Unit
in. inch(es)
ft foot (feet)
yd yard(&)
cu yd cubic yard(s)
mi mile(s)
gal gallon (s)
lb pound(s)
ton short ton(s)
sec second(s)
h hour(s)
day day(s)
gpm gallons per minute
mgd million gallons per day
hp horsepower
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CONVERSION TABLE
Data are reported in this document primarily in SI metric units. Custom-
ary U.S. units are also given in parentheses to make the data fully
accessible to sections of industry which have not yet undertaken metric-
ation. Where the data are reported from (or compared with) earlier
studies which did not use SI metric units, the customary U.S. units are
used as the primary expression of measurement.
SI Metric
1 kilometer (1000 m)
1 meter
1 centimeter (1 x 10" 2 m)
1 nanometer (1 x 10~9 m)
1 square meter
o
1 hectare (10 000 m )
1 cubic meter
1 liter (1 x 10~3 m3)
1 gram
1 milligram
1 kilogram
U.S. Equivalent
0.621
3280.84
1093.61
39.370
3.281
1.094
0.394
mile
feet
yards
inches
feet
yards
inch
3,281 x 10~2 foot
39.370 x 10~9 inch
3.281 x 10~9 foot
1550.16
10.765
square inches
square feet
3.861 x 10" 3 square mile
6.102 x 104 cubic inches
2.642 x 102 gallons (U.S. fluid)
35.314 cubic feet
1.308 cubic yards
0.264
gallon (U.S. fluid)
35.314 x 10~3 cubic foot
2.205 x 10~3 pound (avoirdupois)
2.205 x 10~6 pound (avoirdupois)
2.205 pounds (avoirdupois)
xi
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SI Metric
1 tonne (1000 kg)
1 meter per second
1 cubic meter per second
CONVERSION TABLE (continued)
U.S. Equivalent
1 cubic meter per day
1 cubic meter per meter per day =
1 liter per second
1 watt
1 kilowatt (1000 watts)
1 microsiemen per centimeter
o
2.205 x 10 pounds (avoirdupois)
1.102 short tons
3.281
2.237
35.314
feet per second
miles per hour
cubic feet per
second
2.282 x 10' gallons per day
22.82 million gallons
per day
1.585 x 10^ gallons per minute
35.314 cubic feet per day
2.642 x 102 gallons per day
10.765
cubic feet per foot
per day
35.314 x 10~3 cubic foot per
second
15.850 gallons per minute
2.282 x 10^ gallons per day
1.340 x 10" •* horsepower
1.340 horsepower
1 micromho per
centimeter
xii
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ACKNOWLEDGMENTS
Under subcontract to W. A. Wahler & Associates, Environmental Control
Technology Corporation (ENCOTEC) of Ann Arbor, Michigan provided services
in water quality monitoring and chemical analyses, and Chase-Mogdis, Inc.
of Ann Arbor, Michigan provided services in economic analyses and cost
comparisons. Toy Drilling Company of Distant, Pennsylvania subcontracted
drilling and pumping services.
The personnel of the Lancashire No. 20 mine near Carrolltown, Pennsylvania,
owned and operated by the Barnes and Tucker Company, provided valuable assis-
tance to the study team. Despite the inconveniences imposed upon them by the
study, they were most cooperative in supplying mine data, cost data, and water
quality information, in surveying drill holes, and in providing general assis-
tance underground, including recording flow measurements and collecting water
samples. We are particularly indebted to Mr. George Merritts, Chief Engineer
and Mr. Jerry Sousa, Mining Engineer, for their active support, patience, and
sustained interest. We owe thanks also to Mr. Robert W. Roland, Assistant
Vice President of Finance of the Barnes and Tucker Company, for his cooperation
and for his provision of costing data.
The U. S. Bureau of Mines and various state agencies contributed informa-
tion and advice in a friendly and cooperative manner. The Department of
Environmental Resources and the Bureau of Water Quality of the Commonwealth of
Pennsylvania were of particular help. Valuable guidance was provided by the
staff of the U.S. Environmental Protection Agency, Industrial Environmental
Research Laboratory in Cincinnati, Ohio, in the administration and performance
of the study. Special recognition is due to Mr. S. Jackson Hubbard, Project
Officer.
xiii
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SECTION 1
INTRODUCTION
Acid mine drainage (AMD) is a major pollution problem in the eastern coal
mining regions. In Pennsylvania alone, Klingensmith (1965)1 estimates 3700 km
(2,300 mi) of streams have been adversely affected. Underground mines contri-
bute a large portion of the acid drainage water which constitutes a widespread
pollution problem throughout the eastern coal mining regions. In underground
mines AMD is formed when iron disulfides (e.g. marcasite, pyrite) decompose on
exposure to air, forming a complex series of compounds that are highly soluble
in water. Groundwater inflow to the mine, which may be of excellent or poor
quality when it enters the mine, dissolves these compounds, forming acid water
characteristically high in iron.
Natural groundwater flow patterns are changed by an underground coal mine,
which acts as a large subsurface drain or sink, inducing flow into the mine.
The inflow rates are dependent on the hydraulic characteristics of the sur-
rounding rock units, the degree of fracturing of these units, and the hydraulic
gradients induced by drainage into the mine. Several types of rock can act as
primary aquifers that store and transmit groundwater. In the bituminous coal
region of the eastern United States, interbedded sandstones, shales, silt-
stones, and some limestones associated with the coal strata serve these func-
tions. In particular, sandstone channels may be sources of large inflows where
they are in contact with the mine opening and have a relatively high permea-
bility. Where fracture-dominated flows occurs, all of these rock types can
act as secondary aquifers and provide storage space to recharge fracture zones.
The formation of AMD has been extensively studied in recent years along
with various techniques for water quality control. Water quality control can
be achieved either by treating drainage water outside the mine or by pre-
venting contact of air or water with the iron disulfides. Treatment, usually
by neutralization, is the most common method currently in use, but it is ex-
pensive and not too satisfactory for various reasons. In particular, the
addition of calcium in solution tends to increase total dissolved solids and
hardness of the water and the sludge generated may be difficult to dispose of
in a safe manner. Prevention of air contact with acid-forming minerals—for
example, by using remote controlled mining methods—holds some promise for the
future. Water contact should be controllable by interception of groundwater
inflows to the mine (dewatering).
Interception of groundwater inflows to active coal mines is the subject
of this study, which has involved the construction and operation of a pilot
well dewatering system to collect and analyze both technical and cost data.
The Lancashire No. 20 mine, located near Carrolltown, Pennsylvania, was the
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study site for the pilot dewatering program. This site was selected for
several reasons: the mine inflows are not only easily monitored, but are of
high quality in contrast to the mine drainage, which requires treatment before
it can be discharged; and the owner, the Barnes and Tucker Company, was very
cooperative and interested in the study.
The objectives of the pilot study were (1) to determine the impact of the
dewatering operation on the quantity and quality of inflows in the limited
area of the mine used for study; (2) to evaluate, by extrapolating the results
of this program, the potential effectiveness of the dewatering technique for
the mine as a whole; and (3) to perform an economic evaluation of the tech-
nique. Consideration of several potential sites for the study and review of
information available in current literature, State and Federal agencies, and
industry provided material for technical and economic comparisons.
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SECTION 2
CONCLUSIONS
The following is a summary of the major conclusions reached from the
results of this study.
1. Though several of the rock strata that occur in association with coal
measures can act as aquifers, the inspection of a limited number of mines
during this study suggests that fracture-dominated flow systems are prevalent
where large mine inflows of groundwater are experienced. The larger inflows
are usually localized along fracture zones associated with faulting and/or
jointing, which in turn are often associated with folds.
2. Exploration and pump tests of wells revealed that groundwater flows into
the study area of the mine were concentrated along two fracture zones that
appear to intersect in the area of greatest inflow. These preferred and
rather narrow flow paths were not revealed until after completion of obser-
vation wells and dewatering wells. This late recognition indicated a need
for better exploration techniques to recognize and define the characteristics
of such flow paths, as discussed in Section 3.
3. Baseline monitoring of water quantities established that flows respond
rapidly to wet and dry periods in spite of the depth of the study mine—over
152 m (500 ft)—and thus indicated that fracture zones have hydraulic
connection to the surface. Baseline water quality data indicate that deg-
radation of mine waters occurs gradually as the water is transferred through
the mine; thus the water does not require treatment until it reaches the last
few transfer stations.
4. The quality of water pumped from dewatering wells remained high, permit-
ting direct discharge to Laurel Lick Run. In some respects the water quality
was higher than the water quality generally occurring in the receiving
streams. The intercepted groundwater would be of benefit to stream water
quality because of the higher alkalinity and pH which would add buffering
capacity to the creek. Therefore, large discharges from a full-scale
dewatering system could benefit a stream in this manner, making it less sen-
sitive to degradation by acid drainage.
5. Pilot-scale dewatering, after several trials and interruptions, was
performed continuously for 14 days, with three wells pumping an average of
5.2 l/s (82 gpm). The inflow to the study area was reduced from 6.94 l/s
(110 gpm) to 4.6 l/s (73 gpm), or an average of 34 percent. The average well
effectiveness was initially 45 percent; that is, 45 percent of the 5.2 l/s
(82 gpm) pumped was diverted from the mine inflow. The well-effectiveness
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increased to 55 percent near the end of the test. From these data it was
projected that well-effectiveness would increase to about 80 percent after
120 days of pumping with no recharge. Therefore, dewatering well-effective-
ness for full scale mine dewatering was estimated conservatively to range
between 50 and 80 percent for this mine.
6. The actual range of costs experienced by various mine operators for treat-
ing and disposing of approximately the same quality and quantity of water is
very difficult to define and compare accurately because of differences in
accounting practices used by the mine owners. However, some useful comparisons
can be made, as long as their limitations are recognized, and it is apparent
that the range of costs for any one element of treatment and disposal can be
great. For example, the cost of water treatment alone can vary from less
than IOC/1,000 gal to more than $1.00/1,000 gal, depending on the quality of
water to be treated and partly on the water treatment plant capacity or flow
rate.
For the study mine, costs for water collection and treatment were com-
puted from company records. It was found that the total capital investment
for water collection was $432,000, whereas that for water treatment was
$430,000. The annual operating cost for water collection and treatment is
approximately $739,000. Of this total, approximately $530,000 is collection,
and $209,000 is for treatment. Of the $530,000, approximately 77 percent goes
for water collection and transfer costs, and pumping to the surface accounts
for 23 percent. Based on an average flow rate of 3.1 million gpd the annual
operating costs presented above convert to approximately 46.8<:/l,000 gal
for collection and 18.50/1,000 gal for treatment. Major cost contributors
to the $530,000 annual operating costs are: Power, 37 percent; labor, 36
percent; material, 19 percent; other charges 8 percent. Methods of deter-
mining the above figures are presented in Section 9.
The computed values of treatment plant capital cost ($430,000) and annual
operating expenses for treatment ($209,000) compare reasonably well with
figures from published sources for other plants of similar capacity treating
water with similar flow rates and acidity levels.
7. Individually pumped wells constructed from the surface, as used in this
study, do not appear to be cost-effective to control water quality at the
Lancashire No. 20 mine unless the average well yield can be increased three
to four times the 1.9 t/s (30 gpm) used in the analyses. The cost of well
dewatering at this mine appears to be, on the average, at least twice as
great as present water removal and treatment costs. If the acidity of the
mine water were higher (in the range of 500 to 1500 mg/£) this dewatering
system would appear attractive. Also, if the coal seam were less than about
45 m (150 ft) deep, it appears that dewatering would be cost-effective using
individually pumped wells. In the cost analysis, rather conservative well
yields were used based on the assumption that it would be difficult to locate
wells along fracture zones, where optimum yields could be obtained. If
fracture zones can be located accurately on the surface and penetrated with
wells, then it is quite possible that dewatering with this type of system
could be cost effective. An average well yield of 1.9 -t/s (30 gpm) was
assumed in this analysis but, with wells located only in the more intensely
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fractured rocks, yields could average as high as three to four times this
amount. If an average well yield of 5.7 to 7.6 £/s (90 to 120 gpm) could
be obtained, the system of individually pumped wells would be cost-effective.
8. The individually pumped surface wells used in this study constitute the
only dewatering system that can be used completely in advance of mining. How-
ever, this is the most costly method and requires the most surface facilities.
Wells drilled from the surface to intersect mine openings and drained by grav-
ity into the mine are less costly and do not require permanent surface facil-
ities except for discharge points, but they are limited as to location and
they do require collection of flows into the mine and pumpage to the surface.
Angled drain holes drilled from the mine openings and drained by gravity into
the mine appear to be the least expensive option and are potentially the most
effective means of mine dewatering where they can be used. They can be more
easily varied to meet unanticipated underground conditions, and the system
can be constructed by mine crews using conventional underground drilling
equipment. This system can be used to dewater to a limited degree in advance
of mining and depth of the coal seam has little influence on cost-effective-
ness. Calculated costs for these alternative systems (presented in Section 9)
suggest that dewatering by means of angled drain holes would be at least
marginally cost-effective at the study mine and that this system is worth
further consideration.
9. Cost-effectiveness analyses did not consider indirect benefits of dewater-
ing such as reduction of production losses due to high water inflows and un-
stable roofs. Such production losses due to poor water control can be much
more expensive than the acid drainage problems which the water also creates.
Dewatering might additionally be combined with methane removal, as well as
other improvements that could enable sharing costs. The cost-effectiveness
analysis was not a full cost benefit comparison, which is required to deter-
mine the true economic feasibility of mine dewatering.
io. It was noted that water degrades progressively as it passes through the
mine (item 3 above), and additional monitoring of water transferred through
the mine indicated opportunities for separation of good and poor quality
waters. Therefore, an alternative to dewatering for control of AMD would be
to collect water underground at the point of emergence and discharge it
directly before it becomes degraded. This can be done through bore holes
located near the areas of inflow and/or with pick-up points and pipelines
through the wet areas in the mine. Poor quality water from sources such as
old mine workings can be intercepted and treated separately, reducing the
amount of water to be treated.
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SECTION 3
RECOMMENDATIONS
1. Better exploration techniques are needed to locate and define the extent
of fracture-controlled flow paths. Work has been done and is in progress
using high- and low-level imagery. This should be supplemented with surface
geophysical techniques such as various resistivity surveys and spontaneous
potential streaming surveys. These exploration techniques should be proven
with test wells and observation wells and hydraulic characteristics determined
from pump tests.
2. A pilot dewatering study should be developed for drain holes drilled from
the mine openings to develop better cost- and well-effectiveness data. This
study could probably be best conducted by the mine operator with specialized
help where needed.
3. Water quality sampling within a cross-section of active coal mines should
be conducted to determine the opportunities available for prevention of water
quality degradation as water is transferred through the mines. Also, a
demonstration project could be performed to determine the cost-effectiveness
of separating and selectively treating or discharging waters from various areas
of a mine.
4. Exploration programs for coal mine development should include geohydro-
logic data collection. Expensive exploration holes should serve the dual
purpose of providing groundwater data as well as information on coal seams.
Natural groundwater can be sampled and analyzed for water quality. Holes
should be cased to bedrock and preserved as water level monitoring wells.
Water level contours can provide valuable information on the preferred flow
paths typical of fracture-dominated flow systems.
5. A study is needed to identify the benefits of mine dewatering in addition
to water quality control and prevention of mine flooding. Anticipation or
prevention of large, sudden groundwater inflows could prevent losses of
production time. The improvement in roof stability that could be expected
from dewatering in advance of mining would have direct beneficial effects on
such problems as lost production time and safety hazards. Potentially, such
improvements in operating conditions could bring much larger economic benefits
than would control of AMD.
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SECTION 4
BACKGROUND
2
Various investigators have studied the use of well systems to dewater
underground coal mines. These investigators have identified most of the
technical problems associated with well dewatering systems, and the equipment
and technology for construction and pumping are readily available. In
practice, however, there is little or no experience of well dewatering of
underground coal mines (except for shaft dewatering) that is available to be
drawn upon, and the well dewatering systems that have been published are
idealized and not necessarily compatible with field conditions. It is
questionable whether mathematical techniques are useful in designing well
systems for an entire mine where ground-flow is dominated by fracture perme-
ability, which usually results in preferred and erratic flow paths. If the
individual flow paths can be identified and their hydraulic characteristics
determined, it may be possible to model groundwater conditions mathematically.
However, from a practical standpoint, developing these parameters with exist-
ing technology presents a very difficult and costly problem. Judging from
the experience gained during this study, the condition of fracture-dominated
flow seems to be the rule rather than the exception in wet mines of the
eastern coal regions.
GENERAL HYDROGEOLOGIC CONDITIONS
In the eastern bituminous coal region, coal seams are typically associated
with a series of sandstones and shales (including siltstones, claystones, etc.)
with some limestones. Bedding tends to be tabular with gradual pinching out,
except for sandstone channels which are lenticular in cross-section. Bedding
is generally nearly flat-lying or has gentle dips associated with broad folds
or regional tilting. Low-displacement faults are common in many areas, and
jointing is variable but often associated with axes of gentle anticlines and
synclines. Despite the relatively uniform geology throughout the region, it
is dangerous to make generalizations with regard to groundwater conditions.
In the absence of fracturing, the sandstone, limestones, and coal seams
act as aquifers because their permeability is relatively high compared to
that of the shales, which act as confining beds. Technically, if this solely
were the case, it would be rather easy to determine aquifer constants from
pump tests and mathematically to predict groundwater response to imposed
situations, i.e., dewatering wells. However, fracturing due to jointing and
faulting cuts across bedding planes, locally destroying the confining effect
of relatively impermeable beds, and bedding plane fractures or partings can
create relatively high permeabilities in shales. In zones of relatively
intense fracturing, permeabilities are often several orders of magnitude
-------
higher than in unfractured rocks, resulting in fracture-dominated flow along
fracture planes. This preferred flow creates extremely difficult problems in
predicting groundwater movement and response to induced drainage.
Among the mines inspected during this study, fracture-dominated flow,
which tends to localize inflows, appeared to be prevalent in all cases. A
critical step in coping with this type of flow system is to develop techniques
to locate and define the extent of fracture zones and to develop practical
testing methods to determine their hydraulic properties. Considerable work
has been done with aerial photographs to locate these zones and further work
is in progress. Even if the hydraulic properties of fracture zones remain
unpredictable, definition of these zones will make mine dewatering from wells
much more practical.
Even in fracture-dominated systems, sandstones, limestones, and coal
seams still act as secondary aquifers and probably provide most of the
groundwater storage space. If the intersecting fracture zone is drained by
a mine opening or by wells, water is released from storage. However, the
geometry of the potentiometric gradients becomes complicated because the cone
of depression is elongated along the fracture zone (as shown on Figure 1) and
the driving forces exerted on the secondary aquifers inducing their drainage
is also distorted. Also, drawdowns within the fracture zone may be well
below the base of secondary aquifers and may therefore result in variations
in discharge, as well as in head, with time. Further, the individual fracture
planes in a zone are probably interconnected erratically, so that groundwater
response in one fracture plane may be greatly different from the response of
a parallel fracture plane a few feet away.
Discussion with other investigators and review of literature indicate that
there are areas in the region where fracture-dominated permeability does not
control groundwater transmissions. Sandstones, and apparently sandstone
channel deposits in particular,_can have a wide range in intergranular perme-
ability perhaps ranging from 10 8 to over 10 3 cm/s. Limestone beds also
can be highly permeable if solution cavities are present, and these are
probably the second most common source of high mine inflows after fracture
zones. (Alluvial and glacial deposits in contact with mine openings are
considered a special case.) As indicated previously, these rocks can be
treated readily with conventional hydraulic relationships once the hydraulic
parameters and extent and thickness are determined through exploration and
testing.
HYDROLOGIC EFFECTS OF UNDERGROUND MINING
Under natural conditions, the surface of the zone of saturation tends to
conform in a subdued manner to the topography and surface drainage systems.
Variations occur, depending on the presence of geologic structural features
such as fracture zones and folding or tilting of beds. The variation tends
to become more pronounced with depth, where geologic structure provides the
primary control. An underground coal mine, as it grows, disrupts natural
groundwater flow patterns by acting as a large drain. The flow of groundwater
into the mine results in lowering of potentiometric levels above and around
the mine. Overlying and underlying aquifers will be affected to varying
8
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DRAWDOWN CONE PARALLEL TO FRACTURE ZONE.
FRACTURE ZONK
2. DRAWDOWN CONE ACROSS FRACTURE ZONE.
Figure 1. Schematic asymmetry of drawdowns along a fracture zone,
-------
degrees depending on the leakage characteristics of confining layers or
boundaries and depending on the extent to which aquifers are interconnected
by fracture zones.
Drainage to the mine is increased by fracturing related to subsidence,
which differs somewhat according to the mining method employed. This effect
may not be noticeable with room and pillar mining until pillars are removed
or until long after the mine is abandoned. Longwall mining methods let the
roof collapse in the longwall panel as coal is extracted. With both mining
methods subsidence and fracturing may occur to the surface and the result is
the development of a new hydrogeologic environment over the mine. This is
more apparent with longwall mining because the induced fracturing sometimes
results in large increases in mine inflow until most of the stored water
intercepted by the new fractures is drained. With both mining methods, in-
flows of groundwater increase with expansion of the mined area. It is
generally believed that the increase during the active mine life appears to
be generally greater with the longwall method, although this belief has not
been substantiated.
Drainage into mines can deplete overlying aquifers and, of course, affect
wells and springs, and stream flow also can be intercepted. Often the
heaviest mine inflows are beneath streams, because stream courses are often
structurally controlled by synclines, fractures etc., and the streams can
provide more recharge to fractures than can water stored in secondary
aquifers. Limited observations during this study support this reasoning.
Groundwaters entering active underground coal mines may be of excellent
quality or they may be degraded to various degrees by natural processes or by
migration through adjacent or overlying abandoned mines. Analyses of uncon-
taminated groundwaters associated with coal seams in Illinois and Indiana^
indicated pH values of 7.0 or above and total dissolved solids concentrations
ranging from approximately 1000 to 60 000 mg/£. In the Appalachian region
the quality of uncontaminated groundwater is generally much better, though
there is a general increase in dissolved solids with depth. A representative
list of analyses showing ranges of natural water quality for this region was
not found, but experience during the study suggests that total dissolved solids
for mines 183 m (600 ft) deep or less should often be less than 500 mg/l, with
a near-neutral pH. This better quality is probably due to the fact that water
transmitted through fracture systems circulates more rapidly, tending to flush
more mineralized waters or to preclude their formation.
PROBLEMS ASSOCIATED WITH MINE WATER
The problematic effects of AMD have been extensively described by others
and will not be discussed further here. Likewise, the techniques of treating
AMD are well known by the industry and readily available in published litera-
ture. Neutralization is the most common method of treating AMD.
Most of the direct costs of handling and treating the mine water itself
can be readily identified. However, the indirect costs of any water-related
10
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problems are difficult to separate from operating costs. Loofbourow and
Brittain* outline the problems that are effects of wet conditions in mines:
1. Direct Effects
a. Cost.
b. Failure to handle inflow may interrupt production and
damage the mine, perhaps beyond recovery, perhaps with
loss of life.
2. Indirect Effects in Mine
a. Freezing water in shafts in cold areas.
b. Reduced efficiency of crews and equipment.
c. Added maintenance of equipment, tires, cost or hazard
of using electricity.
d. Washed ground.
e. Flows of hot water heat and humidify ventilating air.
f. Dissolved gas carried into some mines.
g. Reduced stability of rock walls (and roofs).
h. Increased work on accessways and ditches.
i. Interferes with certain explosives.
j. Makes some products unacceptable.
k. Scale in pipes and pumps.
3. Indirect Effects Outside Mine
a. Increased shipping, treatment, handling costs.
b. May pollute surface water (and groundwater).
c. Drawdown may take water from wells and lower quality
of water remaining, or improve it.
d. Drawdown may cause surface subsidence.
e. Drawdown may affect surrounding water users.
Many of the problems cited here relate to the simple presence of water in
the mine, not directly to the quality of the water. However, some or all of
these problems may be positively affected by a solution of the water quality
problem that is based on the prevention of water entering the mine in the first
place. Some of these problems can have much greater impact on mining costs than
that of routine water handling and treatment—for example, interruption of
production due to large, sudden inflows or to unstable roofs. Mine dewatering
has the potential of reducing the magnitude of such problems and the associated
costs. However, these problems vary greatly from mine to mine, as well as
varying within a mine from one area to another and from one time to another.
The true overall cost benefits of dewatering have to be evaluated on a case-by-
case basis. This is best done by the mine engineering staff, with specialized
help were needed, because they are most familiar with the operational con-
straints and conditions of their mine.
11
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CURRENT MINE WATER CONTROL
In shaft and slope mines, water inflows are controlled usually by a
collection system of free drainage, ditches, sumps, and pipelines used to
transfer water (by gravity flow or pumping) to stations from which it is
pumped to the surface. The amount of pumpage used to transfer water to the
final discharge point varies greatly with the gradients in the mine and the
amount of water handled. This system is flexible in that the collection and
transfer stations are expanded as the mine grows and water inflows increase.
At the working faces, water is collected in small sumps either by gravity
drainage or by small, portable sludge pumps. From the initial sump or
collection point, water is transferred through a series of stations that
collect increasing volumes of water until the primary discharge station is
reached where the mine water is pumped to the surface.
The capacity of the system must be large enough to accommodate large,
unanticipated inflows. This is one of the basic weaknesses of the system in
that the capacity has to be greater than normally used, or mine production is
badly interrupted when large flows are encountered. This is especially true
where fracture-dominated flows are encountered. As fracture zones are en-
countered or as new fractures are created by mining, flows of ten to several
hundred liters per second can suddenly enter the mine. Further, these flows
may peak and subside to a fairly constant rate or remain high over long
periods of time, depending on the recharge source and the volume of ground-
water in storage. This places a continuing but erratic strain on the mine
water handling system. As a result, the system has to be changed or re-
established periodically and both production losses and over-capacity have to
be accepted to some degree. Dewatering in advance of mining or in conjunction
with mine advance can help stabilize this situation.
From a water quality standpoint, the current water collection and trans-
fer methods provide,opportunities for degradation in water quality by contact
with decomposed iron disulfides. This probably occurs mostly on the mine
floor, in open ditches, and in open, unlined sumps. As will be demonstrated
later in this report, degradation often occurs gradually as mine water is
transferred from point to point before reaching the surface. The relatively
concentrated inflows might not change significantly in quality if they were
collected in the immediate inflow area and transferred from sump to sump by
pipeline, even if the sumps were unlined and open. In contrast, widely dis-
tributed drips and seeps that are allowed to accumulate slowly on the mine
floor and are collected by sludge sumps probably become acid relatively fast.
If this widespread inflow is mixed with concentrated inflow that far exceeds
it in volume, degradation may not be significant if the water is transferred
to the surface rapidly. It is therefore apparent that opportunities probably
exist for water quality control underground with current mine water handling
systems that could at least reduce treatment costs. This might be accom-
plished by collecting water at transfer stations close to the source of inflow
and pumping either directly to the surface (through boreholes) for discharge
to receiving streams, or through a closed system to prevent further exposure
to acid-forming minerals.
12
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SECTION 5
SITE SELECTION
A major task of the mine dewatering study was to select an active coal
mine in the eastern coal mining region that would be suitable for a pilot-
scale dewatering operation. It was decided during project planning to limit
the search generally to the region north of the Ohio River in Illinois,
Indiana, and Ohio, but to include northern West Virginia and western
Pennsylvania. This area, it was reasoned, would provide a good cross-section
of the eastern bituminous coal mining region. Also, this area would be more
convenient and would therefore facilitate field work by the study team
members based with the Ann Arbor, Michigan firms Environmental Technology
Control Corporation (ENCOTEC) and Chase-Mogdis, Inc. (C-M). The search team
organized by W. A. Wahler & Associates consisted of a hydrogeologist and a
mining engineer (both Wahler personnel), a sanitary engineer (ENCOTEC), and
an economist (C-M). This team established site selection criteria based on
study objectives, schedule, and budget, and divided the selection process into
three phases. The three phases consisted of a literature search, a telephone
survey, and site visits. Through a process of elimination, this three-phase
search resulted in selection of the Lancashire No. 20 Mine, located near
Carrolltown, Pennsylvania and owned by the Barnes and Tucker Company.
SELECTION CRITERIA
The search team prepared lists of criteria to use as general guidelines
in comparing mine sites. These criteria were flexible because it was realized
from the outset that an ideal site probably would not be found. However,
minimum conditions had to be met to result in a successful program. The
conditions listed below were considered as essential criteria:
1. The mine containing the test site must be an active coal mine in the
eastern coal mining region.
2. The mine operator must be cooperative and interested in the program.
3. The mine test area must be overlain by saturated rocks with signif-
icant inflows of groundwater into the mine.
4. Mine inflows must be relatively good quality water, but must degrade
in quality in the mine, and require treatment before discharge to
receiving streams.
5. The test area in the mine must be capable of isolation so that inflows
can be measured and sampled separately from other mine water.
13
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6. The land surface over the test site must be accessible to drilling
and test pumping equipment.
The study was directed toward the active portion of coal mines, although
in many cases it was found that the major sources of water quality problems
were adjacent abandoned underground workings. The degradation of ground water
in abandoned mines prior to its entry into an active mine is a problem con-
sidered to be beyond the scope of this study since it requires a different
approach. Inflows of both degraded and non-degraded waters occur (as is the
case in the Lancashire No. 20 Mine) but they must be identified and handled
separately.
The cooperation of the mine operator was a key site selection criterion.
The operator had to be willing to be subjected to a number of inconveniences
in supplying information, including operational and capital cost data, and in
providing access,to the mine for underground monitoring activities. Without
this cooperation, and in effect, direct participation and interest in the
results of the study, performance of the program would have been impossible.
Saturation of rocks overlying the mine throughout the year was essential
for several reasons. Only saturated groundwater flow can be intercepted
effectively with dewatering wells. Inflows into the mine must be large enough
to create an impact on mining operations by requiring handling and treatment.
Groundwater in storage above and adjacent to the mine had to be sufficient to
provide a fairly constant inflow and dampen the effects of recharge events due
to storms.
Groundwater inflows to the mine must be of reasonably good quality and
must require relatively little or no treatment for discharge into receiving
streams, but after entry into the mine must become degraded to the extent of
requiring treatment for discharge. This criterion was necessary to provide
a contrast for comparison of costs between existing methods of handling mine
water and interception of mine inflows by dewatering.
The capability of isolating a test area of the mine was important because
it was not expected that a significant impact could be made on the total inflow
to the mine, or even a major part of it, with a test operation of limited size.
The test area must be isolated so that flows could be monitored for flow rate
and sampled for water quality.
The last essential criterion was that the land surface over the mine test
area must be accessible to drilling equipment. It was believed that the only
practical way to develop test data would be to drill wells from the surface and
intercept groundwater inflows by pumping. Land ownership, access, cultural
features, and topography were all matters of concern.
Another major consideration was the depth of the mine at the test area.
The optimum depth range for construction and operation of dewatering wells
was believed to be 45 to 120 m (150 to 400 ft). It was felt that shallower
test areas might be subject to fluctuations of inflow and water levels due to
recharge from storms. Deeper wells would be much more difficult and costly
to construct and pump.
14
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Other factors that were of secondary importance, but that were considered,
included coal seam thickness; mining method; and the availability of hydro-
geologic, water quality, and cost data. The coal seam thickness and mining
method were considered mainly with respect to working room and accessibility
of the underground test area. The availability of information would influence
both the effort expended on data collection and exploration and also the
quality of the work product.
SELECTION PROCESS
The selection process was divided into three phases consisting of a
literature search, a telephone survey, and site visits. The Keystone Index
of coal mines and state lists of mines were used as the primary reference
to develop an initial candidate list. The mines selected from the Keystone
Index were limited geographically to Illinois, Indiana, Ohio, Pennslyvania and
northern West Virginia. Further, the mines were limited to slope or shaft
mines because it was reasoned that drift mines would be more likely to have
more seasonal water inflow problems and would tend to be located in more
rugged topography. Also, mines deeper than 120 to 150 m (400 to 500 ft) were
avoided. Review of other literature was not fruitful in identifying potential
candidate sites, but it did provide general information on the magnitude and
type of mine water problems.
With the information available, a telephone survey was performed. Both
mine operators and state and federal agencies were contacted throughout the
survey area. This survey provided clues to general mining areas and to
specific mines or mine operators where mine water was a problem. It soon be-
came apparent that the most suitable sites were probably in southeastern Ohio
and western Pennsylvania. The candidate list was narrowed to eleven mines as
a result of this survey, and appointments were made with mine operators to
discuss the program and with state and federal agencies having localized
knowledge of specific mine water conditions.
Discussions with mine operators reduced the candidate list to nine mines
in Ohio and Pennsylvania. Formal permission was then requested both to in-
spect site conditions and to use the mine in question for the pilot dewatering
study if it proved to be the most suitable site. Negative responses were
received from the owners of four mines, which reduced the number of candidate
mines to five, all in western Pennsylvania.
During the spring and early summer of 1976, the five candidate mines were
visited and inspected in detail. The mines visited were the Vesta No. 4,
Shannopin No. 2, and Lancashire No. 24D mines owned by the Jones & Laughlin
Steel Corporation; the Lancashire No. 20 mine owned by the Barnes & Tucker
Company; and the Rushton mine owned by the Rushton Mining Company. The mine
operators were very helpful and cooperative at all five mines. Each site was
inspected at least once, and additional surveys were performed at three mines
before a final choice was made. After the first inspection, the Vesta No. 4
and the Shannopin No. 2 mines were eliminated for technical reasons. Addition-
al visits had to be made at the three remaining sites to collect water samples
for analysis and other detailed data for the final comparison. The suitability
of each mine for the proposed study is summarized in the following description.
15
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Vesta No. 4 Mine
The Vesta No. 4 mine located near California, Pennsylvania is a shaft
mine using room and pillar extraction of coal from the Pittsburgh seam of the
Monongahela Group. This mine was eliminated as a test site because of the
suspected nature of the water inflows. The coal seam crops out in a stream
channel; inaccessible mine workings at a shallow depth may have partially
subsided, permitting stream inflow through fractures. The inflow was esti-
mated to be over 126 L/s (2,000 gpm) where it could be monitored. The chances
of intercepting a significant part of this inflow with wells was uncertain at
best. The risk of developing inconclusive data appeared to be too high.
Shannopin No. 2 Mine
The Shannopin No. 2 mine, located near the Vesta No. 4 mine, is a shaft
and drift mine using room and pillar extraction of coal from the Pittsburgh
seam. This mine was eliminated because the mine drainage water does not
require treatment for discharge to receiving streams. Hydrogeologic con-
ditions appeared amenable to dewatering; however, it was obvious that the
cost-effectiveness of dewatering would be negative from the standpoint of
water quality because the water did not require treatment.
Lancashire No. 24D Mine
The 24D is classified as a slope mine using a longwall method of extrac-
tion of coal from the Lower Freeport ("D") seam, which is about 1.0 to 1.2 m
(40 to 48 in.) thick. This mine, located near Barnesboro, Pennsylvania, was
eliminated because the water pumped from the mine does not require treatment
before it is discharged into the receiving stream. Initially it was thought
that the quality of the mine water was degraded enough to consider the site.
However, after the samples had been collected and analyzed, the results in-
dicated otherwise, as shown on Table 1. Otherwise, the mine met all of the
requirements for the test site. The coal seam is 70 to 90 m (250 to 300 ft)
beneath the surface and the major inflows occur from a combination of
fractures and a sandstone channel along a syncline.
TABLE 1. WATER QUALITY - LANCASHIRE NO. 24D MINE
Total Total Total
Alkalinity Acidity Sulfate Iron
Sample pH mg/£ as CaCO^ mg/l as CaCO me/I mg/l
Mine Inflow 7.45 138 — 110 0.038
Mine Discharge 7.92 78 — 210 0.48
16
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Rushton Mine
The Rushton mine located near Philipsburg, Pennsylvania is a shaft mine
using room and pillar extraction of coal from the Clarion ("A") seam. This
mine was eliminated because of poor quality water inflow to the mine. Initial
inspection of this mine indicated that it had most of the prerequisite condi-
tions needed for the test site. It appeared that one heading in a fault
block structure with a possible sandstone channel would serve as an ideal test
area. It had an added advantage of occurring only 60 to 75 m (200 to 250 ft)
beneath the surface. However, the water inflows were sampled and analyses
revealed that the water quality was degraded before entering the active mine.
The source of this poor quality water is apparently some abandoned workings
in the Lower Kittanning ("B") coal seam, which is about 25 m (80 ft) above.
Groundwater above the Lower Kittanning seam is. of good quality, as shown on
Table 2.
TABLE 2. WATER QUALITY - RUSHTON MINE
Sample3 pH
Total
Alkalinity
as CaC03
Total
Acidity
mg/£ as CaC03
Specific
Conductance
yS/cm
Sulfate
mg/£
Total
Iron
mg/l
R-l
R-2
R-3
R-4
R-5
R-6
R-7
4.43
4.45
3.07
3.04
6.80 4
3.67
6.30 3
437
909
428
540
-
42
1270
2100
1220
1680
1500
270
50
1060
2040
1020
1320
1150
96
21
180
364
102
163
0.97
4.02
0.04
Location of Water Samples:
R-l
R-2
R-3 Mine discharge water before
treatment
Roof, East Main
Roof, first left North Main
Mine discharge water before
treatment
R-5 Treated Discharge water
R-6 River water above Rushton
Treatment Plant discharge
R-7 Shallow well water above
"B" seam
17
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Lancashire No. 20 Mine
The No. 20 mine is located within a few miles of the No. 24D mine. It is
also a slope mine utilizing the longwall mining method to extract coal from
the Lower Kittaning seam, which is about 1.5 m (60 in.) thick. Table 3 shows
that the quality of mine inflow water is very good and, if intercepted, could
be discharged directly into a stream without treatment. This table also
shows that the quality of the mine water deteriorates as it is transferred
through the mine to the F-l or F-14 sumps, where it is pumped to the surface.
The water then requires treatment before discharge to the receiving stream.
Several inspections revealed that a small area of the mine with reasonable
inflows—6 to 10 £/s (100 to 150 gpm)—could be readily isolated for mon-
itoring of flows and water quality. Mine inflows were transmitted by
fractures in a thin shale bed overlying the coal seam. It was thought that a
thick sandstone bed overlying the shale would act as an aquifer which could
be penetrated by wells to intercept mine inflow. The principal problem with
this site was that the test area of the mine is 150 to 170 m (500 to 550 ft)
beneath the surface and wells had to be placed on a moderately steep sidehill.
Also, surface rights are not owned by the mine, and therefore access rights
would have to be acquired from the landowner. However, after comparisons with
the other available sites, the Lancashire No. 20 mine was selected as the best
alternative for the pilot dewatering operation.
TABLE 3. WATER QUALITY - LANCASHIRE NO. 20 MINE
Sample
PH
Total
Alkalinity
mg/£ as CaCO
Total
Acidity
mR/t as CaCO
Total
Sulfate Iron
Mine Inflow 8.04
Mine Discharge
before Treatment
F-l Sump 3.03
F-14 Sump 6.27
Surface Water Above
Treatment Plant
Discharge 7.42
115
105
15
26
41
880
610
44
1.59
50
12
1.40
The Barnes and Tucker No. 20 Mine Treatment Plant treats the combined
F-l and F-14 flows and by NPDES permit must produce an effluent with a pH
between 6.50 and 9.0, with a total iron concentration not to exceed 4.0 mg/£.
18
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DEVELOPMENT OF A PILOT DEWATERING PLAN
In conjunction with final site selection, a program plan was developed
for the pilot dewatering operation. The program plan was discussed by the
project team and formulated in late July, 1976. The program consisted of the
following elements:
Baseline data collection
Construction of a test well and observation wells
Initial pump testing and pilot dewatering system design
Construction of the dewatering system
Operation of the pilot dewatering system
Schedule and costs
The basic plan was to develop wells adjacent to the underground test area
that would develop a lateral barrier to mine inflow and reduce heads in the
sandstone above the test area. It was decided to maintain a distance of 30 m
(100 ft) from the mine face for safety reasons except for one observation well
partially penetrating the overlying sandstone. Since the mine was known to
have a methane problem, it was judged to be too risky to penetrate the coal
seam any closer without disrupting mining operations.
The dewatering plan was predicated on the assumption that the sandstone
bed over the coal seam would act as an aquifer. This turned out not to be the
case. Fracture permeability dominated groundwater flow at this site and the
fractures cut across bedding planes, resulting in unpredictable hydraulic
characteristics of the rock units. Instead of being able to design wells based
on pump test data from a single well, the dewatering system had to be developed
almost on a trial-and-error basis because hydraulic responses could not be
predicted for even a few meters. As a result, each dewatering well had to be
developed and tested separately, and even these tests turned out to be deceptive.
19
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SECTION 6
SITE CONDITIONS
DESCRIPTION
Mine
The mine selected for study In this Investigation, the Barnes and Tucker
Company's Lancashire No. 20, is an active mine located near Carrolltown,
Cambria County, Pennsylvania (Figure 2). The Lancashire No. 20 is a moderately
deep slope mine, and the Lower Kittanning ("B") coal of the Allegheny group
is the only seam mined. The coal seam averages 1.5 m (60 in.)*in thickness
and is located at depths ranging from 150 to 170 m (500 to 550 ft) in the
vicinity of the study area. The "B" coal is a metalurgical grade, low sulfur,
low volatile coal.
The longwall method of mining is used. Coal is cut with an Anderson
Mavor double drum ranging shear, and the roof is supported with a Douty 10-ton
shield with belf-advancing chocks. Coal is transported by 91 cm (36 in.) and
107 cm (42 in.) conveyor belts that exit from the mine through a slope shaft.
Driving of the mains and entries for development of the longwall panel is
accomplished with continuous miners. During this latter operation, coal is
removed by shuttle cars and unloaded on to the conveyer belts. The mine is^
normally operated in three shifts and has a daily capacity of 3200 t (3,500
ton) of cleaned coal.
The mine area lies within the drainage basin of the Susquehanna River,
whose flow ultimately reaches the Atlantic Ocean. Surface runoff from the
part of the mine area selected for the pilot dewatering program drains directly
into Laurel Lick Run, which lies downslope from the test area. Laurel Lick
Run, with a drainage area of approximately 2331 ha (9 sq mi), flows into
Chest Creek (near Bradley Junction), which in turn flows into the West Branch
of the Susquehanna River. Laurel Lick Run and Chest Creek are two of the
streams in the area which have not been severely polluted, and which support
trout and other aquatic life. In the past Chest Creek has been used as a
source of municipal water and could be used in this manner in the future. Be-
cause of the high quality of water in this watershed, treated mine drainage
from the Lancashire No. 20 mine is discharged into the West Branch Susquehanna
and not into Laurel Lick Run, which is closer.
Mine Water Conditions
The Lancashire No. 20 mine, along with many other mines in Cambria County,
has continually had problems with groundwater inflow. This drainage into the
20
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"* s ': / /•"' y . r
. *; v • W
• . - , . rs ?•
••. '. i - . - •- ' i J
,-^j
-v.,,:
,
••"•----.-•• • •••
,•;••
:---^\,
, /"' V
. •; H
!
I' ;• '
——i i
L i v:
<• A
. ' X
<
'
'""
'
•,^.-,i, -
•,^.-,i, -
<' | ••• ? N^-.-
$g\
^^
KEY
•MINE PROPERTY BOUNDARY
-EASTERN WORKING FACE
SCALES
0
1 KILOMETER
1 MILE
CONTOUR INTERVAL 20 FEET
Figure 2. Location map: Lancashire No. 20 mine and pilot well dewatering site.
21
-------
mine has had a significant impact on working conditions and on the overall
cost of mining. To control this inflow, the mine has constructed a water
handling system consisting of pumping stations that transfer water through a
series of pipelines and sumps.
The source of water inflow to the mine is from groundwater in overlying
and surrounding rocks. Water inflows to the mine occur primarily through the
roof and exterior faces and water is transmitted predominantly by fractures.
There is probably also some inflow through the mine floor, but it appears to
be relatively minor in comparison. The mine floor is on an underclay
(clay-shale) that restricts upward seepage of water.
Although minor drips and seeps are scattered throughout the mine, the
larger inflows are localized, probably along zones of more intense fracturing.
Fractures and fracture zones are developed originally by geologic processes
and later by fracturing associated with subsidence of rocks overlying the mine
openings. The natural fractures are encountered while driving main haulage
ways, and while establishing entries and cross cuts during the development of
longwall panels for mining. Unstable roofs are often associated with areas of
relatively high water inflow.
Inflows into the longwall panels and into abandoned areas of the mine are
influenced both by natural fractures and subsidence fractures. The longwall
panels are 150 to 170 m (500 to 550 ft) wide and several thousand meters long.
Normally, all of the coal is extracted in a continuous operation and the roof
support is moved forward with the coal cutting machine, allowing the roof be-
hind the support to collapse. Entries previously driven alongside the panel
partially dewater the panel area, but new fractures associated with roof
collapse, along with pre-existing fractures, promote more drainage into the
mine. New fractures may extend to the surface, resulting in drainage of
groundwater that was not tapped previously. These mechanisms can result in
inflows of water that are large enough to cause handling problems and that
have hydrostatic heads sufficient to affect the caving characteristics of the
roof—both of which can affect production. Similarly, additional fracturing
can be caused by the slow collapse of abandoned or inactive openings in which
roof supports have not been maintained.
It appears that approximately 40 percent of the water enters through
abandoned parts of the mine and 60 percent enters directly into the more recent
workings. Inflows are affected by seasonal conditions and it is even possible
to recognize recharge from individual rainstorms. There is a tendency for
high inflows to occur beneath stream channels, which probably tend to follow
fracture zones along parts of their course. Wells and springs overlying the
mine may be affected or drained completely, especially where subsidence has
occurred.
Waiter Removal and Treatment
Water in the mine is collected and transferred through a series of sumps
and conduits and eventually pumped to the surface. Initially water is col-
lected by- small sludge pumps or allowed to drain by gravity into sumps. From
there it is pumped from sump to sump through pipelines, as shown schematically
22
-------
on Figure 3. Water is pumped to the surface from the F-l and F-14 sumps and
the pipelines are combined to a single flow at the treatment plant, which has
a design capacity of 19 000 nar/d (5 mgd). Flow to the plant is controlled by
pump operators underground, according to water levels in the sumps. Logs are
kept for each pump and are used to determine flow rates, based on the pump
curves for the two sumps. Figure 4 is a graph of average daily flows to the
treatment plant. The pattern of these flows tends to follow the dry and wet
periods for this region. The yearly average flow for the period July 1976
through June 1977 is 13 000 m3/d (3.4 mgd). The data after June 1977 are not
reflective of normal flows, due to the flooding on July 19-20 and cleanup
operations extending well into August 1977.
At the treatment plant, in a chemical mix tank (see Figure 5), lime in
slurry form is added to neutralize the water and to raise the pH to a suf-
ficient level to cause iron precipitation. The lime is added manually, and
quantitites are adjusted according to data from daily observations and weekly
chemical analysis of effluent. Suspended solids and iron precipitate are
then removed in settling ponds, and the effluent from the ponds is piped to a
company-owned facility in another watershed for discharge into the West
Branch Susquehanna River. Effluent from the treatment plant is monitored
weekly for pH, alkalinity/acidity, and total iron. According to data provided
by mine personnel, the average levels for the period June 1976 to August 1977
were:
Average Standard Deviation
pH 8.7 0.7
Alkalinity (mg/£ as CaCO™) 60 20
Total Iron (mg/£) 2.1 1.4
Since the mine owners buy lime in bulk to serve all treatment and process-
ing needs at more than one plant, records of quantities used in individual
situations were limited. However, plant personnel made an intensive effort
over several weeks to define usage specifically for this treatment plant, and
according to the results of this study the average usage was 1360 kg (3,000
lb) of 95 percent CaO (calcium oxide) per day.
Underground jtudy Area
The underground test site herein referred to as the "Main G study area"
was located near the northeasternmost limit of workings along Main G heading,
or (as commonly referred to in the mine) the left side of Main G, near
section G-14. The study area is bounded by the haulage way on the south; by
the edge of underground working on the north and east; and by the G-ll entry
on the west. Figure 6 presents a map of this portion of the mine.
This area was selected because it presented relatively high groundwater
Inflows estimated at 6 to 10 £/s (100 to 150 gpm) and because it could easily
be separated from other mine drainage and monitored. Mining in the immediate
area had been temporarily halted because of unstable roof conditions and high
groundwater inflows, so advancement of the workings would not hamper the
progress of the project.
23
-------
L-2
SUMP
1
•DRAINAGE COLLECTED FROM MAIN G STUDY AREA
G-11
SUMP
G-10
SUMP
6-9
SUMP
F-20
SUMP
G-7 SUMP
MAINH
SUMP
F-13
SUMP
F-17 SUMP
to
G-5SUMP
MAIN F SUMP
OLD WORKINGS
DRAINAGE
SLANT SUMP
K-12
SUMP
K-7
BELT
SUMP
SUMP
F-14
SUMP
OLD WORKINGS DRAINAGE
F-l
SUMP
MAIN C
SUMP
OUT
DRIFT
Figure 3.
TO TREATMENT
Water transport system.
-------
NJ
20
16
14
12
S 10
JUL.
UNSEASONABLY HIGH
FLOW DUE TO PUMPING
OUT MINE FOLLOW INfi
JULY 19/20. 1977 FLOOD
! ' ' i J L
SEPT.
1976
NOV.
JAN.
TIME (Month)
MAR.
MAY
1877
5.0
4.5
4.0
3.5
3.0
2.5
a
to
JUL.
Figure 4. Average daily flows to treatment plant,
-------
F-14 SUMP
F-1 SUMP
LIME ADDITION
CHEMICAL MIX
TANK
SETTLING POND
SETTLING POND
•*• EFFLUENT
Figure 5. Treatment plant - schematic layout.
26
-------
*>z =*=
MODERATE
INFLOW
GRAVITY FLOW
SMALL POND
L-4
HEAVY INFLOW
70-75 GPM
LIGHT INFLOW
uTpobT
i TT^
CEMENT '
BLOCK DAM
K E Y
DIRECTION OF
UNCCNTAINED FLOW
-*-DIRECTiON OF
PIPED FLOW
25
100
SCALE
I)
•V.
METERS
n
•%.
FEET
25
100
SECTION
6-11
B.C. - BUTT CLEAT
F.C. - FACE CLEAT
'15
NO. OF
>FRACTURE
I MEASUREMENTS
COURTESY OF A. IANNACCH IONE ,
U.S. BUREAU OF MINES.
FRACTURE FREQUENCY DIAGRAM
Figure 6. Main G study area.
27
-------
Water enters the study area of the mine from the roof and from the coal
seam. The roof is a gray to dark gray-brown shale-siltstone, which is frac-
tured to intensely fractured. Slickensides and polished surfaces are common
on much of the rock, and water drains freely from these fractures. Fractures
in the roof rock are common throughout this area of the mine. Water inflow,
however, tends to be more localized in areal extent. The shale-siltstone
roof is about seven feet thick and is overlain by a massive sandstone.
Originally, it was thought that this sandstone was acting as an aquifer and
that drainage from it was the source of groundwater inflow to the study area.
However, drilling and pump testing indicated that this unit is relatively
tight except along fractures.
The area of greatest mine inflow is along the L-4 heading near the
northern face (see Figure 6) where most of it is collected behind a cement
block dam. Overflow from the L-4 pond and any additional drainage from the
left (north) side 'of Main G heading are collected in a low area of the mine
floor, forming a pool located at the east end of the L-2 heading. Prior to
establishment of the underground monitoring system, the L-2 pool also col-
lected drainage from the right (south) side of the Main G heading. Water
from both the L-4 pond and L-2 pool are then transferred to the L-2 sump,
which is formed by a cement block dam and is also located in the L-2 heading.
Gravity flow is utilized to drain the L-4 pond, whereas pumping is required
to remove water from the L-2 pool. Water collected in the L-2 sump includes
all of the mine inflow to the Main G study area. From L-2 sump, water is then
pumped to the G-ll sump where it is combined with water from adjacent areas
and transferred through the mine and eventually to the surface, as shown
schematically on Figure 7.
INVESTIGATIVE ACTIVITIES
The investigative activities! performed for this project had a dual purpose:
first, to establish the geologic and hydrogeologic conditions for use in the
design, construction, and operation of the pilot dewatering program; and
second, to develop baseline data on inflows and water quality. The results
of the hydrogeologic aspects of the investigation appear later in this section.
The pilot dewatering program is presented subsequently in Section 7. The
investigative activities included:
• Development and construction of underground monitoring facilities.
• Reconnaissance geologic mapping.
• Drilling and construction of observation and dewatering wells,
with logging of drill holes to provide hydrogeologic data.
• Pump testing of dewatering wells.
• Monitoring of underground and surface waters to develop baseline
data.
Investigative activities began in September 1976, with the establishment of
the underground monitoring facilities and surface geologic mapping. Drilling
and pump testing operations followed these and continued into December, when
extremely difficult weather conditions slowed down progress. Due to the
severity of the weather, field activities had to be temporarily demobilized
in late January. Operations resumed in May and all drilling and testing
28
-------
-STEVENS RECORDER
L-2 SUMP
WATER METER
GRAVITY
FLOW
(L-4 POND)
—J> Q(L-
SAMPLING
STATION
PUMPED
2 POOL)
STUDY AREA
G-11 SUMP
OTHER INFLOWS
G - SECTION
G-7 SUMP
—Q
OTHER INFLOWS
G - SECTION
to
\e>
SLANT SUMP
F-1 OR
F-14 SUMP
TO SURFACE AND TREATMENT PLANT
OTHER MINE
INFLOWS
OTHER MINE
INFLOWS
Figure 7. Mine water flow path and monitoring system - Main G study area to treatment plant.
-------
activities were completed in June 1977. Monitoring activities continued at
least on a limited scale without a break until pilot dewatering operations
were initiated in July 1977.
Underground Monitoring
The underground monitoring facilities were designed to obtain background
data on quality and quantity of mine inflow, and to provide a means for de-
tecting changes in these during the pilot-scale dewatering. To monitor mine
inflow, it was necessary to isolate the underground study areas from surround-
ing mine drainage. This was accomplished by diverting all flow from the right
(south) side of the Main G heading to the G-ll sump located to the west of the
test site. Mine drainage from the study area was then collected in the L-2
sump described previously.
After this diversion was completed, only water inflows from the study area
were collected and pumped from the L-2 sump. Figure 7 schematically shows the
water flow path out of the mine and the monitoring system established for this
study.
Monitoring the rate of mine inflow was accomplished using a 7.8 cm
(3 in.) Rockwell flow meter attached to the discharge line of the pump drain-
ing the L-2 sump. The water level in the pond was monitored in conjunction
with this to record periods of pumping and non-pumping. Water-level monitoring
was accomplished using a Stevens water-level recorder. Mine inflow was then
calculated using data obtained from both of these measurements.
Samples for water quality analysis were obtained from a sampling box (see
Figure 8) built near the end of the discharge line from the L-2 sump to the
G-ll sump. Samples were collected using a Protech pressure-operated composite
sampler. No rock particle sediments accumulated in the box during the
monitoring period. Additional samples for water quality analysis were ob-
tained from inflows from the roof near the L-4 pond. These were "grab"
samples collected directly from the mine roof. All monitoring and sampling
equipment used was inspected by state and federal mine inspectors prior to
its installation in the mine.
The flow meter was monitored and water samples from the Protech sampler
were collected by mining personnel. Flow meter readings were taken each
shift, or three times daily, during the routine inspection of the pump at the
G-ll sump. Twenty-four-hour composite water samples from the sampler were
collected only during the pilot dewatering operation.
The Stevens water-level recorder was monitored by personnel from the study
team. The water-level recorder was set on an eight-day time cycle, and chart
paper was changed weekly. During each mine visit, the underground study area
was inspected to insure the continued separation of mine drainage and the
continued collection of flow meter readings. Grab samples of inflow near the
L-4 pond were also periodically collected during these visits.
30
-------
FLOW PUMPED FROM
MAIN 6 STUDY AREA
76.2mm (3") ROCKWELL TURBINE
WATER METER
PLATES TO REDUCE
TURBULENCE
y
SAMPLING LINES TO
PROTECH WATER SAMPLER
i I
*
_D
•OVERFLOW
LINE TO G-1
SUMP
SAMPLE CHAMBER
(NOT TO SCALE)
m DIMENSIONS
H- 0.6m(2 it)
W- 0.8m(2.5 ft)
L-0.9m(3 ft)
Figure 8. Sampling box at G-ll sump (schematic).
31
-------
The monitoring system as described was in operation from September 1976
to July 1977. During the night of July 19-20 the mine was flooded as a con-
sequence of surface runoff into the shaft following a severe thunder storm
(which flooded Johnstown and other communities in the vicinity), and mining
operations were temporarily halted until the mine could be dewatered.
Following storm water removal and cleanup at the mine, the underground
water transport scheme was reorganized and isolation of the left side as pre-
viously maintained was no longer possible. The new system required temporarily
directing all flow from the left side away from L-2 sump and toward L-2 pool.
This was accomplished by plugging the gravity line draining L-4 pond and al-
lowing the water to flow by gravity to the L-2 pool, where all inflow from
the left side was now collected. (L-2 pool had been previously used to
collect overflow from L-4 pool and other additional drainage from the left
side which did not flow directly into L-2 sump.) Monitoring and sampling
were performed as described earlier; however, both instruments were now
located at L-2 sump near the end of the L-2 pool discharge line. No change
was noted in the chemistry of the monitored flow as a result of rearranging
the monitoring devices. The revised monitoring system and water flow path
is shown schematically on Figure 9. This new system had two disadvantages.
First, it allowed more lag time between any changes in mine inflow and actual
measurements. Secondly, recirculation of L-2 pool water was possible, since
overflow from L-2 sump would flow by gravity into L-2 pool. If this occurred,
a higher inflow rate would be calculated if corrections were not made. To
correct for times of overflow, the Stevens recorder was again monitored at the
L-2 sump.
Geologic Mapping
The ground surface overlying the underground test site was geologically
mapped in September 1976, prior to the start of exploratory drilling. The
purpose of the mapping was to determine whether surface evidence existed for
geologic control of mine inflow and also to aid in locating future dewatering
wells. Low-elevation aerial photos were reviewed in the performance of this
task. Generally the terrain proved not to be amenable to mapping and this
part of the investigation contributed only minor information regarding pos-
sible well locations. However, it was possible to identify the general
geologic features of the site area. The reconnaissance geologic map (Figure
11) is presented with the geology discussion later in this section.
Well Construction and Pump Testing
Well Construction - A total of seven wells was drilled from the surface
to obtain subsurface geologic data and for use in the well dewatering program.
Well locations are shown on Figure 10, and drill logs are presented at the
end of this section on Figures 16-22. All drilling was done with air-rotary
rigs, initially with a single compressor Davy rig and later with a larger,
dual compressor Schramm rig. Drilling and all subcontracted field work was
performed by Toy Drilling of Distant, Pennsylvania.
Initially three wells were drilled. Locations of these wells were de-
termined primarily by areas of high inflow in the underground test site, and
32
-------
RIGHT SIDE
INFLOW
Ui
L-2 SUMP
G-7 SUMP
SLANT SUMP
F-1 OR
F-14 SUMP
TO SURFACE AND
TREATMENT PUNT
-STEVENS WATER
LEVEL RECORDER
SAMPLING STATION
WATER METER
ALL LEFT SIDE
INFLOW PUMPED
GRAVITY FLOW
Figure 9. Revised mine water flow path and monitoring system - Main G study area to treatment plant,
-------
::::x'>:::::::::::::^::::::::::::::::::::::::::::::::^^
x::::::ox^:::::>:::::::::::::::^:^::::::;:::::::::^:v::::::::::::::::::^^
25
100 50
SCALE
D
J-
METERS
n
FEET
25
100
Figure 10. Location map: Dewatering and observation wells.
34
-------
by surface topography and accessibility. Test well P-l, 20 cm (8 in.) in
diameter, was located approximately 30 m (100 ft) horizontally from the outer-
most mine workings in L-4, and drilled to the base of the "B" coal seam. Two
15 cm (6 in.) diameter observation wells, OB-1 and OB-2, were drilled in con-
junction with P-l. These holes were located 9 m (30 ft) and 37 m (120 ft),
respectively, from the mine workings. Termination depths were 15 cm (50 ft)
above the mine roof for OB-1, and the base of the "B" coal seam for OB-2.
During the drilling of each of these wells, drill cuttings were inspected and
identified and a drill log prepared. This and comparable data derived from
holes drilled later were used to construct a stratigraphic column (Figure 12,
present on page 40 with geology discussion). Also, water encountered in and
air-lifted from the holes as drilling progressed was logged to help determine
the locations of water-bearing zones. The data from these observations, along
with pump test data, were used as the basis for determination of the site
hydrogeology.
Following the drilling of these three wells, P-l was pump tested. Analysis
of pump test data indicated that pumping had a significant impact on mine in-
flow, and two additional test wells and two observation wells were located
and constructed. Three of the additional wells were over 150 m (500 ft) deep
and were drilled to or within 15 m (50 ft) of the "B" coal seam. The fourth,
a shallow observation well, was drilled to 58 m (190 ft). All additional
wells were drilled and logged in a manner similar to the initial three wells.
Following pump testing of P-l and prior to any additional testing, all six
of the deep wells were partially cased to seal off upper water-bearing strata.
A Hookwall packer was attached to the base of the casing: 15 cm (6 in.)
diameter casing for test wells and 10 cm (4 in.) diameter casing for the
observation wells. The packers were firmly seated in a massive sandstone unit
in an attempt to prevent falling water in the wells. The shallow observation
well (OB-3) was left uncased to detect drainage effects on shallow rock units.
Well construction details are shown on drill logs at the end of Section 7,
Figures 16-22 (pages 51-57).
Pump Testing Dewatering Wells - Pump testing was accomplished using a
3.7 kW (5 hp) submersible pump with a gasoline generator as an electrical
power source. Constant discharge tests were run for variable lengths of
time and water levels were monitored continuously throughout the pumping and
recovery periods. Stevens continuous float-operated water-level recorders
were installed on the observation wells to monitor changes in water level.
Measurement was made through small diameter PVC tubing which was affixed to
the pump discharge line and which extended to within 3m (10 ft) of the pump.
Pumpage rates in the pumping well were determined by measuring the length of
time necessary to fill a 20 I (5 gal) bucket. Pumping was on the order of
1.5 to 1.9 £/s (25 to 30 gpm), so this was easily accomplished.
P-l was initially pump tested in November 1976. OB-1 and OB-2 were used
as observation wells in this test. All three of these wells were uncased at
that time. Pumping in P-l was at a rate of 1.9 £/s (30 gpm) and continuous
for 19 hours. Drawdown measurements were recorded at both the observation
wells. Because of problems with the water-level indicator in the pumping well,
°o drawdown measurements were recorded in P-l. P-l was again pump tested in
35
-------
May 1977. All six of the surrounding wells were completed at that time and
were used as observation wells. Partial casing was installed in P-l and all
the observation wells except OB-3. Testing was for 2.5 hours at a rate of
1.9 £/s (30 gpm). Drawdown and recovery measurements were recorded in the
pumping well and in all of the observation wells.
P-2 was pump tested in January 1977. OB-1, OB-2, OB-3 and P-l were used
as observation wells. Pumping initially was at 1.9 tfs (30 gpm); however,
the well went dry in less than 30 minutes. The pumping rate was then re-
duced to 0.5 £/s (7.9 gpm); the well again went dry. The pumping rate was
then reduced to 0.15 £/s (2.3 gpm). Pumping was continuous for one hour.
Drawdown and recovery measurements were recorded in the pumping well; however,
no effect was seen in any of the surrounding observation wells.
P-3 was pump tested in June 1977. All six of the surrounding wells were
used as observation wells. All the wells were partially cased except OB-3.
Pumping was at 1.46 £/s (23.1 gpm) and continuous for 8.5 hours. Drawdown
and recovery measurements were recorded in both the pumping well and the ob-
servation wells.
Selection of Pumps and Generator - On the basis of pump testing and mine
inflow data, pumps were selected for the pilot well dewatering scheme. The
3.7 kW (5 hp) pump used previously in testing was installed in Well P-2,
which produced minimal pumpage during the testing. A 10 cm (4 in.) diameter
5.6 kW (7.5 hp) pump was installed in Well P-l and, because a second 10 cm
(4 in.) was not available, a 15 cm (6 in.) diameter 5.6 kW (7.5 hp) pump was
installed in Well P-3. Installation of the latter pump required the removal
of casing from P-3. No significant change in water level occurred following
the removal of casing. A diesel Caterpillar generator was used as an
electrical source for these pumps.
Water Quality ^Sampling and Analytical Test Procedures
Samples for water quality analysis were collected at the Main G study area
and on the surface periodically throughout the performance of this study.
Samples were initially collected to obtain background quality data on mine in-
flow and on the creek adjacent to the proposed dewatering wells. Later,
samples were collected to monitor the effects of pump testing and pilot de-
watering. Water discharged from the dewatering wells was also sampled to
insure that well discharges met the requirements of the Commonwealth of
Pennsylvania and to detect any changes in water quality.
In addition to sampling at the underground study area, mine drainage was
periodically sampled from one of the two mine drainage exit sumps (F-14) prior
to its entrance into the treatment facility. A sample of mine drainage was
also collected for the other exit sump (F-l) and at several of the other
underground sumps in an effort to trace water quality in the mine. This
latter sampling was not thought to be required for the pilot dewatering study.
However, it was relevant to the consideration of alternative methods for
water quality control.
36
-------
A field laboratory was set up near the mine site to provide the capability
of analyzing samples shortly after their collection. The field laboratory
also provided the ability to obtain certain water quality results quickly
throughout the study program. Supplies and equipment in the field facility
were provided for performing the following analyses: pH, alkalinity, acidity,
sulfate, ferrous iron, and specific conductance.
All analyses were performed using methodology approved by the U.S.
Environmental Protection Agency (EPA), Quality Assurance Branch. Alkalinity
was analyzed by titration with standard suo-furic acid to an electrometric end
point of pH 4.5. Acidity was determined by titration with standard sodium
hydroxide to an electrometric end point of pH 8.2. The turbide-metric tech-
nique using barium chloride was utilized for sulfate determinations. Ferrous
iron was analyzed by using the phenanthroline method in conjunction with a
spectrophotometer set at 510 nm and providing a light path of 2 cm. Other
analyses, such as total dissolved solids, total iron, calcium, and magnesium
were performed at ENCOTEC's main laboratories in Ann Arbor, Michigan. Samples
collected for these analyses were preserved according to EPA procedures and
shipped to the laboratory for analysis. All metal analyses were performed
using atomic absorption spectrophotometry by the flame method.
Abandonment
Following the completion of all field activities, including the pilot well
dewatering, the casing in all the wells was pulled and the wells were cement
and concrete grouted to the surface. Abandonment procedures were in accordance
with the standards set by the Commonwealth of Pennsylvania, Department of
Environmental Resources.
HYDROGEOLOGY
.Surface Geology
Unconsolidated materials cover most of the surface of the area and include
Quaternary alluvium, colluvium, and sedentary soil. Sedentary soil and
colluvium make up the greater percentage of this cover and typically consist
of silty sand to silty clay. It forms a soil cover over bedrock units that
varies widely in thickness. Quaternary alluvial material consists chiefly of
sandy silt and sand. This alluvium is restricted to the flood plain and
channel of Laurel Lick Run.
All rocks exposed on the surface are of Pennsylvanian age and belong to
the Glenshaw Formation of the Conemaugh Group. Typically they consist of
thinly to thickly bedded sandstone and shale. Sandstones are generally
quartzose and/or argillaceous, fine to very fine grained, moderately hard to
very hard, moderately weathered, and iron oxide stained. Shales are gray
to gray brown, moderately hard, and moderately to severely weathered. Local
residents stated that a thin coal seam outcrops in the immediate site area;
however, it was not found in the field. The outcrops of sandstone and shale
are relatively few in number and are generally covered by soil units. For
this reason, these rock units (with one exception) were not differentiated in
Capping and they were grouped as one unit with the colluvium and sedentary soil.
37
-------
The only mappable bedrock unit was the Saltsburg Sandstone, a member of
the Glenshaw Formation of the Conemaugh Group. Outcrops of this unit are gen-
erally small; however, a 6 to 7.5 m (20 to 25 ft) vertical section is exposed
along a railroad cut in the central portion of the site. Previous geologic
mapping by Campbell and others (1913)5 mapped this unit as the Buffalo
Sandstone. Data cited in their report, along with other published geologic
literature on the area, and drilling data included in this report were used
to identify this rock unit. Figure 11 shows the areal geology in the immedi-
ate vicinity of the study area.
Published geologic maps and the geologic mapping of the area indicate that
the strata are generally flat-lying, with a slight dip to the southeast. No
faults or other linear features were identified in either the field investi-
gation, or in the review of low-altitude photos of the area.
Seeps, springs, and other hydrologic features were also noted during the
surface geologic mapping and are identified on the geologic map. Springs were
few in number and not associated with any given horizon or topographic
elevation. Only flows which were continuous year round were designated as
springs, and flows for these were generally greater than 0.3 t/s (5 gpm).
Seeps, on the other hand, were quite common. Flow is usually less than 0.05
to 0.1 tfs (1 to 2 gpm), occurs intermittently, and is associated with wet
periods. Seeps are commonly found in the topographic lows or swales or in
areas with a break in slope.
Subsurface Geology
A stratigraphic column of units encountered in the drilling program is
shown on Figure 12. This is a generalized column and represents a composite
of all information obtained during the drilling phase.
All rocks encountered in the subsurface are from the Pennsylvanian period
and are sedimentary in origin. Lithologically they are sandstone, shale,
siltstone, limestone, coal, and clay. Figure 13 is a cross-section of the
site based on well data. Correlation between wells was accomplished using
several of the coal beds which occur throughout the section. Since drill
cuttings rather than cores (for cost reasons) were the source of log data,
not every stratigraphic unit was identified in each well, especially some of
the thinner beds. Where this was the case, correlation lines have been dashed
or queried.
Rocks encountered in the subsurface can be divided into two stratigraphic
groups: the Allegheny and the Conemaugh. The Allegheny Group, the older of
the two, can be further divided into three or five formations. Flint (1965)
subdivided it into three formations (oldest to youngest): the Clarion, the.,
Kittanning, and the Freeport Formations. On the other hand, Edmunds (1969)
subdivides it into five formations (oldest to youngest): the Clearfield Creek,
the Millstone Run, the Mineral Springs, the Laurel Run, and the Glen Richey.
The stratigraphic column (Figure 12) follows Flint's division of the group.
The lowermost formation in both Flint's and Edmunds' subdivisions was only
partially penetrated—1.5 m (5 ft)—in the borings and is not shown. Maximum
thickness encountered in the wells for the Allegheny Group is 76 m (250 ft).
38
-------
GEOLOGIC UNITS
RECENT ALLUVIUM; FLUVIAL DEPOSITED SAND AND SILT.
CONEMAUGH GROUP, MASKED BY COLLUVIAL MATERIAL.
100
25 METERS
3
100 FEET
CONTOUR INTERVAL 5 FEET
SYMBOLS
SALTSBURG SANDSTONE; (MEMBER OF GLENSHAW FORMATION OF CONEMAUGH GROUP)
LIGHT YELLOW BROWN TO GRAY BROWN, VERY FINE GRAINED.
QUARTZOSE. MODERATELY HARD TO VERY HARD; LOCALLY
CROSS-BEDDED; TYPICALLY BREAKS INTO TABULAR FRAGMENTS.
SPRING
— GEOLOGIC CONTACT- DASHED WHERE APPROXIMATELY LOCATED; SMALLER DASHES
WHERE PROJECTED BY DRILL HOLE OR TOPOGRAPHIC DATA.
Figure 11. Areal geology.
39
-------
SYSTEM
GROUP
FORMATION
DESCRIPTION
COLUIVIUll ~ 0-7-9 "1 (0-26-); SLOPE WASH OEJRIS; TYPICALLY, NODEHATE BROW* TO REDDISH 6ROWN SIITY-CLAYEY SAND HMD
SALTSBURE SANDSTONE - 0-25.0 m (0-821); LIGHT mm 10 MIT BROIH, VERY FIHE TO NED I UN GRAINED PEI GRAVEL LOCALLY
PRESENT, QUARTOS. ARGILLACEAOUS. MODERATELY HARD TO YERY HARO; WEATHERED NEAR SURFACE. SONE UC»L NODERATELY WEATHERED
OTHERWISE FRESHj FE» INTEK8EDS OF SHALE AND SILTSTONE; EROSIOKAL LOWER CONTACT.
SHALE. LIMESTONE - M. 6-18. 6 n (48-61' ); LIGHT BLUISH GRAY LINESTQHE AND UNET SHALE BOUNDED ST UPPER AND LOWER
SHALE UIIITS. UPPER SHALE: DIRK GRAY. NOOERAIELY HARD. LOWER SHALE: GRAY, IITH SOW R£0 TO REDDISH 8ROWN MEAD BASE
(MEYERSDALE RED BED?); NOOESATELY HARO; INTERBEOS OF SIL1STONE AND SANDSTONE IN BOTH SHALE UNITS.
BUFFALO SANOSTONE-SiLTSONE - 2.4-8.8 m (8-291); 6RAY TO GREENISH GRAY, VERY PIKE TO FINE GRAINED NICACEOUS
ARGILLACEOUS. NOOERATELY HARO TO HARD.
- 10.7-15.8 HI (35-52'); GRAY TO GRAY BROUN. SANOY, MODERATELY HIRO; SONE INTEH8EDOED SILTSTONE.
BRUSH CREEK MAjjINE SHALE - 6.1-8.5 m (20-28'); DARK CRAY. NOOERATELY HARD; DLAd ORGANIC MATERIAL CONMON' CONTAINS
A THIN LINESTONE BED IS en (6") THICK NEAR BASE.
BRUSH CREEK COAL - 0.3 m (12").
CQPISTH SANDSTONE - 5.8-10.4 m (19-34'); GRAY 10 LIGHT BflOVH, YERf PINE GRAINED, OUARTOSE VERY HARD- CONTAINS A
THIN LINESTONE BED IN UPPER HALF OF UNIT.
- 11.3-15.2 HI (37-50'); GRAY TO LICHT BDCIN. NODERATELY HARD; INTER8EDS OF SILTSTOHE AND VERT FINE GRAINED
SANDSTONE; RED SHALE NEAR NIDDLE Of UNIT (UHOMING RED BED?).
XAHONING SANDSTONE - 4.9-7.4 m (is-261); UEHT GRAY TO GRAY BRWN. VERY FINE GRAINED. DUARTOSE, YERY HARD LOCAL
INTERBEDS OF GRAY TO GRAY BROIH SNALE, CRAOATIONAL CONTACT IITH OYERLTING SHALE.
UPPER fREEPORT LIMESTONE - 4.9-5.8 m (16-131); DAM GRAY SHALE IN LATESAL CONTACT IITH GRAY LINESTONE AND LINEY
SHALE OF VARIABLE THICKNESS; EROSIONAL UPPER CONTACT.
INTERBEDDED SHA|,E, SILTSTONE AND SANDSTONF - 13.7-15.8 n, (45-52' ); GRAY TO CRAY BROWN ANO GREENISH GRAY YERt
FINE TO FINE MA IKED. ARSILLACEOUS. •IUCEOUS. NOCESATELY HARD TO VERY HARD; CONTAINS POSSIBLE THIN COAL SEAN <|5 c» IB")
THICK 3.0 m (10') I60»t BASE.
LOWER FREEPORT ("D") COAL - 0.3-1.2 m (12-48").
n,' -- - «•". «"T
FINE GRAINED. ARGILLACEOUS. NOOERATELY HARD TO VERY HARO; POSSIBLY CONTAINS 2 THIN COAL SEANS
-------
SOOn
(1900')-
550»-
(I800T
(17001)-
500 1>
(16001)
(1500' )
450ir
d«00')
400n-
.
KSl
P-4
.-_--
-
MM
MM
NOTE: SEE Fl
BORE H
P-3
..
MMMMMI
J
MM
MM
MM
mm
— -~~~~^
§
-
•
c
u
j
I
BRUSH CREEK COAL *
SASE OF CONEMAUGH GROUP
i
^ o
LOIER FREEPORI "0" CO»L -S |
UPPER KiriAKNING "C PRIiE" COAL ;
KIDDIE KITTANNING "C" COAl 1
=
LHER KinANNING "B" COAL :
GURE 10 FOR
OLE LOCATIONS.
;
-.
'
i-AJ^i
^^
mmm
P7=^
•••i
MMMMMMMM-
mm
.
_
»- 1
OB-3
^^
EAST N
A' B
P-3
^
P_"iLl_T
r^T^
-•—
MMMMMMMl
I
.-
-:
: _,
SOUTH
B'
'""
^
•w
•— ™
••w
(1300')
550u
(IBOO'J
-cnoo1)
-500m
-(1600')
-(15001)
-450m
-(1400')
-4DOB
Figure 13. Site cross-section.
-------
The Conemaugh Group overlies the Allegheny and can be divided into two
formations: The Glenshaw and the Casselman. The Casselman Formation lies
stratigraphically higher than the Glenshaw and was not encountered on the
surface nor in the subsurface. Maximum thickness encountered for the Glenshaw
Formation is 100 m (328 ft). Names of members for both the Allegheny and the
Conemaugh Groups, along with relative thicknesses, are indicated on the
stratigraphic column. Names are taken from published literature and are based
on distances from known marker beds.
The Allegheny Group is of local and regional economic importance because
of the included coal seams. Of the major seams regionally mined, four of the
five were encountered in the drilling. From the oldest to youngest, they
were: Lower Kittanning or "B" coal, Middle Kittanning or "C" coal, Upper
Kittanning or "C prime" coal, and the Lower Freeport or "D" coal. All these
coals consist of one or more seams with a variable amount of shale or boney
coal separating them. The Upper Freeport or "E" coal was the only regionally
important coal not identified or notgof sufficient thickness to be recorded
in any of the borings. Flint (1965) states that it is locally absent in
Somerset County because of an erosional unconformity, and this may be a pos-
sible explanation for its absence in the present study area also. Coal is
only of local importance in this section of the Conemaugh Group and only one
seam was encountered in the borings. This was the Brush Creek coal.
Structurally the area is generally flat-lying with a gentle dip to the
east-southeast (see Figure 13). Strata lie subparallel to one another except
for the lowermost coal seam, the Lower Kittanning (a member of the Kittanning
Formation of the Allegheny Group), which dips to the east at a slightly greater
angle. Minor flexuring is present, and no fault features were identified by
correlation of geologic units.
Groundwater Conditions
Attempts were made to identify the principal water-bearing zones as they
were encountered during air-rotary drilling by logging the amount of water
air-lifted as the holes advanced in conjunction with logging of rock types.
Each water-bearing zone encountered should generally increase the amount of
water air-lifted as long as circulation is maintained. It was found generally
there was poor correlation between rock types and the amount of groundwater
encountered. There appears to be some perching of groundwater in the
Saltsburg Sandstone and possibly in the underlying unidentified limestone.
Below this apparent perched zone, the bedded rock units could not be separated
into distinct aquifers or zones with any degree of consistency. In particular,
the various sandstone beds encountered did not indicate higher permeabilities
than even the shales. There were some indications that coal seams and the
Upper Freeport Limestone are relatively permeable, but not with consistency.
Also, there were indications that the thin shale bed above the Lower Kittanning
("B") coal which forms the mine roof in the Main G study area, is relatively
permeable. This information, although not conclusive in itself, led to the
suspicion that fractures controlled the groundwater flow regime.
42
-------
After the first test well and two observation wells were completed, pump
tests were run. The well (P-l) was left open hole for the first test; later,
a blank casing was installed to a depth of 65.5 m (215 ft) with a packer at
the bottom of the casing to seal the upper part of the well, and another pump
test was performed. The transmissivities and storativities calculated from
observation well data were essentially the same for both tests. Well P-l was
tested at a rate of 1.9 £/s (30 gpm) for 19 hours with a maximum drawdown of
40.5 m (133 ft). The drawdown response in the two observation wells was
erratic with respect to time and distance from the pumped well. Transmissiv-
ities calculated from log-log time-drawdown plots from observation well data
were inconsistent with time-drawdown and recovery data from the pumped well.
Observation well data indicated transmissivities more than an order of mag-
nitude higher than data from the pumped well. Distance drawdown plots provided
closer agreement with pumped well data, but only when each observation well
was plotted separately with respect to the drawdown in the pumped well. This
was because drawdown response was very asymmetric and the cone of depression
around the pumped well was elongated along fractures, possibly in several
directions.
This pump test data, along with drilling information and observation of
inflows to the study area underground, all reinforced the opinion that flow
paths were controlled by fracture zones. Also, it was observed that there
was very rapid response to heavy rainfall both in water levels in wells and
in inflows to the mine, indicating hydraulic connection of fracture zones
with the surface. This presented a problem in locating and spacing dewatering
wells because zones of fracturing were not recognizable. The only evidence
for concentration of fracturing was the localization of water inflows to the
mine in the Main G study area. On the surface, the prominent joint sets
strike to the northwest and northeast. Therefore, it was decided to construct
the remaining wells to the north of the underground study area along an east-
west alignment as shown on Figure 10. Since pump tests of P-l indicated a
substantial effect on the monitored inflow, the new wells were placed at
approximately the same distance—30 m (100 ft)—from the northern mine face.
Drilling of these wells indicated variable yields and it was decided to
pump test each well individually. Well P-2 yielded only about 0.06 l/s (1 gpm)
at the maximum available drawdown and was abandoned for use as a dewatering
well. Well P-3 was pumped at 1.46 £/s (23.1 gpm) for 8.5 hours with a draw-
down of 14.2 m (46.7 ft). Well P-4 was not tested because originally it was
designed as an observation well. However, drilling information indicated a
good potential yield, and it was used as a dewatering well during the pilot
dewatering operation.
Using time-drawdown data from the pumped wells (P-l and P-3) and distance
drawdown data, transmissivities were calculated that ranged between 2.5 and
3,7 m3/m-d (200 to 300 gpd/ft). This was later confirmed using data de-
veloped from the pilot dewatering operation and applying a graphical method
developed by Theis and Conover (1963)9 to determine the percentage of pumped
well water being diverted from a drain. Since the percentage of pumped water
being diverted from the drain (mine inflow) was known, along with the period of
pumping and distance from the drain, the value of S/T was determined graphically
(where T = transmissivity and S = storativity). Since the groundwater appears
43
-------
to be unconfined, a storage constant of 0.03 to 0.06 was estimated (storage
constants could not be calculated from pump tests). Solving for transmissiv-
ity yielded approximate values ranging from 1.2 to 3.7 m3/m-d (100 to 300
gpd/ft), which agrees well with pump test data.
Water level contours were constructed with the limited well data as shown
on Figure 14. These contours are estimates because of poor control along the
mine face and north of the line of wells, but they illustrate the probable
flow pattern into the mine study area prior to the pilot dewatering operation
in September 1977. As constructed, they indicate two fracture zones draining
into the mine study area. These fracture zones strike in about the same
direction as the predominant joint sets in the area and they appear to inter-
sect in the area of the highest inflow to the mine. The wider of the two
zones may be the one/ striking to the northwest. It appears that all of the
pumping wells missed the most permeable parts of these zones. In retrospect,
it appears that OB-2 should have been used as a dewatering well and P-3 should
have been located about 18 m (60 ft) to the,east. This illustrates the need
for a practical method and approach for detecting fracture zones to guide
drilling operations.
Hydraulic interconnection between fracture zones and even within a fracture
zone across fracture planes is variable at best. For example, when P-l was
pump tested after all the wells were constructed, much larger responses to
pumping (drawdowns) were recorded in wells P-3 and P-4 than in OB-1 and OB-2,
which were much closer. When P-3 was pump tested, a much larger response
was recorded in OB-2 than in either P-l or P-4, which were both much closer.
Also, when P-2 was pump tested, there was no response in any of the other
wells. From this, one could speculate in retrospect that during the dewatering
operation very little of the water flowing into the mine along the northeast-
trending fracture zone was intercepted. In fact, flows may have increased due
to steeper gradients induced by reductions in flow along the northwest fracture
zone.
In conclusion, data strongly indicate that groundwater drains into the
Main G study area along narrow preferred flow paths. These flow paths are
controlled by fractures and are directionally oriented. Permeabilities
across these flow paths may be orders of magnitude lower than permeabilities
along the fracture planes. The bedded rock units probably act as secondary
aquifers to various degrees and provide some hydraulic interconnection between
fractures, as well as providing groundwater storage space. The fracture zones
appear to consist mainly of steeply dipping fractures that intersect the sur-
face. This permits rapid recharge from direct precipitation and from streams
or ponds at the surface. The secondary aquifers also release ground water
from storage to the fracture zones. The groundwater flow regime is therefore
quite complex and difficult to model with mathematical techniques because of
the severe boundary effects and directional variations in permeability. In
this fracture-dominated flow system, groundwater response to dewatering wells
is nearly impossible to predict with confidence unless the fracture zones are
defined and their hydraulic properties determined. The amount of exploration
work required to do this may be impractical unless the fracture zones can at
least be defined by methods such as aerial photography combined with surface
44
-------
520 (1706)
490 (1607)
W^mm^^
ft^^^i
Illlllillllililllllllllilll;
i-mM§imim9mi§
:mm^mmm^mmmm
i*-
p
y
$£*.•[
illlll;lii-i
SCALE
25
490 (1607)
100
FEET
WATER LEVEL CONTOURS, ELEVATIONS
IN METERS AND (FEET)
Figure 14. Well locations and water level contours 9/15/77,
45
-------
geophysical techniques such as resistivity or electric self potential surveys
to supplement test wells. More field-oriented research is needed to determine
the practical application of these and other techniques to mine dewatering.
BASELINE DATA
Baseline data was collected to: (1) establish water flow data into and
out of the Main G study area and the mine; and (2) develop background water
quality data in the mine, from test wells, and from Laurel Lick Run. In
general, water inflows to the mine and background water qualities were about
as expected when the site was selected. Inflows to the Main G study area
were at the lower end of the range originally estimated—6.3 to 9.5 £/s
(100 to 150 gpm). Degradation in water quality was found to be slower than
anticipated as the water was transported through the mine. As will be dis-
cussed further in later sections, it was found that the most seriously de-
graded water comes from abandoned or inactive portions of the mines. This
was not investigated directly but it can easily be deduced from the data
developed. Background water qualities for groundwater from the study area
and from Laurel Lick Run were about as expected.
Underground Flow
Plow measurements at the underground monitoring station began in the late
part of September, 1976. Figure 15 represents typical background flow data
for the total water inflow occurring in the Main G study area of the mine.
The data shown are average weekly values and the standard deviation for each
7-day period. These data are based on average daily flows derived from in-
dividual flow meter readings, and have been adjusted for changes in L-2 sump
water levels where appropriate.
The average inflow rate for the entire period of September 26 through
-June 5, 1976 was 6.69 Us (106 gpm) with a standard deviation of ±0.57 £/s
(9 gpm). As can be seen on Figure 15, the water inflow did vary somewhat
over time, but such changes were generally gradual in nature. Some fluctua-
tion in measured flow did occur from day to day but usually followed a given
increasing or decreasing trend over several days. Changes occurred in con-
junction with wet (heavy rain or snowmelt) periods and dry periods. It did
not appear that major fluctuations in flow occurred within a given 24-hour
period. Variations in the data on a very short term basis (less than one day)
were more a function of variations within the pump transferring the water
through the meter than of actual large changes in water inflows into the mine.
Background flow monitoring in the Main G study area continued until the first
pilot dewatering period began in July, 1977. The data developed are similar
to those presented on Figure 15.
The total mine discharge to the treatment plant, averaging 13 000 m /d
(3.4 mgd), was discussed previously, and the average monthly flows are shown
on Figure 15. The primary source of data regarding the mine water treatment
plant flow was the pumping logs of the pumps in the F-l and F-14 sumps. These
logs contained information on the hours each pump ran, and flow to the treat-
ment plant was then determined based on the hours each pump ran and on the
respective pump curves for each pump. Records of all pumps connected to the
46
-------
10.0
150
9.0
8.0
7.0
S.O
130
*
•
ta
110
100
STANDARD DEVIATION
5.0
• -
I.I. I . 1 . I . I . 1 . I
Bl
9/26 10/10 10/24 11/7 11/21 '12/5 12/19 1/2 1/16 1/30 2/13 2/27 3/13 3/27 4/10 4/74 5/8 5.
'22
1976
1977
TIME (Week of)
Figure 15. Average mine inflow - Main G study area 9/26/76 to 6/5/77.
-------
treatment plant were available and checked periodically by the study team.
The F-14'line carrying water to the treatment plant had one location near
the surface where a water meter could be installed. The meter was installed
in this line from November 1976 until April 1977. These data were obtained
in order to augment and check the pump curve estimates. A similar meter could
not be installed on the F-l line due to its location. Hence, the accuracy of
Using pumping logs and pump curve characteristics to predict flow rates could
not be evaluated for this line.
Water Quality
Water samples were collected and analyzed to provide baseline water qual-
ity data in the Main G study area and at selected points through the mine be-
fore the mine water is treated and discharged. The results of this sampling
are presented in Table 4, Baseline Water Quality Data. Samples of both roof
inflow and L-2 sump water were collected to compare quality of the ground
water entering the mine with the quality of the water at the primary under-
ground monitoring station. In establishing the L-2 monitoring station, it
was assumed that water quality would not deteriorate significantly in the
Main G study area before it is transferred to other parts of the mine. Com-
parison of the results of this sampling for these two locations shows that
the waters-were generally similar and that the assumption was valid. Roof
samples vere discontinued after November 1976. The L-2 sampling station
> .utilized the Protech sampler (see Figure 8) and mine personnel collected
. .sample's except when this device was periodically serviced.
. .A.series of samples was collected at sumps to trace the quality of mine
. water after it leaves the Main G study area to the treatment plant on the
surface. Single samples were obtained in October 1976 from the G-7 sump,
'slant _$ump,*" F-l and F-14 discharge lines. Additional samples were collected
from'.the F-14 discharge line later in the study. For safety reasons, our team
members did not have access to the F-l line sampling point and the single
sample was obtained by mine personnel. Results of the analysis of the F-l
sample are as follows:
pH 3.03
Acidity (mg/£ as CaCO.) 105
Sulfate (mg/£) 880
Total Iron (mg/£) 50
The analyses of the samples from the G-7 and the slant sump illustrate the
deterioration in quality as the mine water is transported from area to area
and finally discharged to the treatment plant. Water quality is still good
at the G-7 sump with respect to iron, sulfates and pH. At the slant sump,
after mixing with mine water from other areas including abandoned workings,
the quality is poor and would require treatment before it could be discharged
to a receiving stream. Water quality deteriorates further while passing
through the F-14 and F-l sumps and before reaching the treatment plant.
48
-------
The water samples from Laurel Lick Run were collected at Station A,
located upstream from the pilot dewatering area. The alkalinity levels were
generally very low in this creek, indicating a weakly buffered system. On
most occasions the pH remained above 6.5, but on one day (October 6, 1976)
the pH dropped to 3.2. The variation in other parameters was also considerable
and may reflect the influence of coal waste deposits upstream, along with
variations in flow.
One sample of groundwater was obtained from Well P-l during the initial
pump testing. This sample was analyzed to confirm that the quality of water
Pumped from the dewatering wells would be similar to the roof water in the
Main G study area and that direct discharge to Laurel Lick Run would be
acceptable. Analysis of this sample (see Table 4) confirms that the ground-
water quality is similar to mine water at the inflow point and suitable for
direct discharge to Laurel Lick Run.
49
-------
TABLE 4. BASELINE WATER QUALITY DATA
Location
Main G
Study .
Area-
Roof
Water
Under-
ground
L-2
Station
Laurel
Lick
Run-
Station
A
Treat-
ment
Plant
Influent
Line F-14
G-7 Sump
Date
1976
5/21
9/23
10/5
10/7
11/22
11/24
9/23
10/5
10/6
10/7
11/22
11/23
11/24
11/24
6/-
10/5
10/6
10/7
11/22
11/23
11-
10/7
11/22
11/23
10/7
Slant Sump 10/7
Well P-l
11/30
This parameter
Sample
Type
Grab
Grab
Grab
Grab
Grab
Grab
Grab
Grab
Comp.
Comp.
Grab
Comp.
Comp.
Grab
Grab
Grab
Grab
Grab
Grab
Grab
Grab
Crab
Grab
Grab
Grab
Grab
Grab
Alkalinity Acidity
mg/£ mg/£
Specific Total Total
Conductance Dissolved ferrous Total Suspended
)lS/cm Sulfate Solids Calcium Magnesium Iron Iron Solids Potassium Sodium
^H CaCO, CaCO,
-------
-SYSTEM
-6ROUP
'-FORMATION OEPTH
KU KTERS FEET DESCRIPTION
CONSTRUCTION
ae
MJ
,
'
GLE.NSHAI FGR««TION
DN FREE PORT FORK*! ION
V
i
STEEl
CIS IIC
1.0.
(12")!
STEEL
EISIIO
I5.2W
I.D.
(ILL
P1CIEI
IS. Zen
I.D.
(B">
L
%
25
HOLE DIlKTil
!!.!t«(IW 4")
50
i
3 65.5 •
(315')
100-
HOLE BIIBTEI
150-
1S1.I "
(SIS')
^3j
•
mmm
----'-
— --
—
••*•
IBM
wmtm
mm
1— ^ ^ —
0-15.2 IP - 5ALTSBUB6 SANOSTONt: LI6HT Mm TO GUI Illim. IEIT FIIE 10 FIIE GUIIEO. PEI CIITEL LOCHLT
(0-50') PIESEIT: OUIITDSE. IIEIUICEDUS. 10DEIIIELT HUD TO IEII HIIO; IE1THEIEO IEII StIIIICI IOC1LLI
SOIE lODEUTEll IEITIEIED. IIIEIIISE FIESH; FEI IHTEH1E8S OF SHILf 110 SILTSTOIE.
15.2-31.7 m - SH>Li, LIHESTOHE: CUT LIltSIOUE IND SLICHTLT HEODIS« mm tlHEI SHILE IOUIDEI IT UPPEI 110
LOIEI SHILE: GUT TO DAII GUI IIMI. IITH SOIE REDDISH BUOIN NEII USE (lOERSBIli lit IEDSI)
nOEUTELT HIIO; IITEIIEOS OF SIITSIOIE 1KB SAIDSIOKE; EICDIIDTEIED IITER AT 31.7 m (104').
-100
(1(14-117' )
35.7-49.1 II - SHUt: GUT 10 fill ISM Sill 110 OIK BUI. IODEIITELT HIID 10 TEM HUD- (ITEMED! OF SILTSTOIE
(IH-ISr 1 110 SIHOSTOIE.
-150
49.1-57.3 1" - BRUSH CREF.H MARINE SHALE; OKI BUI. LOCILLI SHOT. IOOEUTELT HUB; ILICI OIGIIIC IITIIIIL
(I6I-I861 ) CmOl; COIIIHS A IHII LIKSIOIE IED < 15 ci (!") THICK.
57.3-57.9 • (188-190' ) - BHUSH CHEEK COAL.
-200 57.9-47. 1 n - CORINTH SAKDSTONt: (10 RETUII OF CUTTIMS>.
(190-220')
67.1-80.2 «l - SHALE: till TO Dill Sill, lira SOIE IEODISH (IHOIIIB IEO IEO?). IODEIIIELI HIIO- IIIEItEDS OF
(220-263') miSTOHE.
-Z50
80. 2-88.1 n - IAHONIN6 SANOSTONi : GUI 10 GUT linn no LIGHT »«W«. IEIT FIIE TO FIIE GUIIEO guilTOSE
(253-289') HIIO ID >EII H1IO; LOCIL IITEIIEOS Of SHILE 410 SILTSIOIE.
88.1-93.0 « - UPPER FREEPOUT LUESTONt, SHALE; BUT UIESTOIE OIEILIII IT Dill GUI SHlLf IITH <«F
300 (239-305') IEIT FIIF GUIIEO SIKJSIDII .
93.0-IOt.5 n - IK7ERBEOOEO SHALE, SUTStOMt AW SAHOSTONE; GUI TO GUT tloil 110 CIEEIISH £UI IEIT
(305-3561) FIK TO FIM 6IIIKO. UOILUtEttlS. IICACEOVS, HDfUlELI HIID 10 IEII Hill- COITIIIS PDSSIILE
IHII COIL SEIIf IS » («") THICK. O.I • (]') IID«E USE.
IOB.5-109.1 « (356-3581) - 101ER FREEfOBT "0" COAL: IICIEIU II FLW IT 101 < . ,««.,
109.1-120.1 n - INIER8EOOEO SHALE. SILTSTONE AND SANDSTOHE- GUI TO CUT noil l«o CIEEIISH GUI
(358-3941) tEll FIK TO Fllf. tllllEO, IIGILLICEOUS. IOOE«IIELI HIID 10 IEIT HIID- IN1I C1IL SEIl'
< l.S en (3") THICK IEII IIDDLE OF Ulll.
120 1-121.0 m (394-3971) - UPPER KITTAIHilNG "C PRIM" COAL.
too
121.0-132.3 n - UPPER IQHTHINCTOH SANOSIOME: CUT. tEll FIIE TO FIIE CIIIIEO. WHOSE, »E«T »IIO-
(397-434') FEI SHALE IIIEIIEOS.
450 132.3-142.6 m - SJAEJ; BUI TO Dill GUI; FEf SILTSIOIE IITEIIEDS; Till COIL SHI IT TOP.
(434-468 ')
14? 6-143.3 m (468-470') - UBOlt KITTANNING "C" COAL.
143.3-160.9 n - LOlEfl I08THINCTON SANDSTONE;
-------
X
X
p
•
.
•
GLENSHAI FORMATION
1
"
i
•
iROur
'ORMATIQN DEPTH
HU «TEHS
CONSTRUCTION
~l
STf.EL
IS INC
I.D.
(12")
STEEL
is me
5.2c«
I.D.
HOOl-
KILL
PICIEI
5.2m
I.D.
I
I
I
(13. 3')
50-
89.2 a
(221')
75-
KOLE 01 METED
100
125
150
• Iffl I •
:'-.•
2^-
-v*-~
mtmrn
.: •," •
^
'•-:^
=5^
••ill
mmm
EET DESCRIPTION
0 ELEVATION 580.8m (1905. 6' )
0-25.0 m - SALTSBURG SANDSTONE: LIGHT BROW TO GRIT BROWN. TERT FIKE TO FINE CRIKEO. dUMTOSE, IH6ILLICEOUS.
(0-8}') IOOERITELT HIRD To'lERT HMD; IEITHERED NE1I SUIFICE. SDK LOCIL lOOEIITELT 1EITKEIEO. OTHEIIISE
5|J FRESH; FEI INTSRBEDS OF SHILE MO SILTSTONE; EROSIONIL LOttl CONTICT; IITER ENCOUNTEIEO IT 12.0 •
<74'>.
25 0-39 6 01 - SHALE LIMESTONE; GRIT LINESTONE INO LIMET SHILE BOUNDED BT UPPER IND LOIEI SHILE INO SILTSTONE
(82-130') UNITS. UPPER Nl'UDIIL: INTERIEOOED SHILE INO SILTSTONE; GRIT TO SRIT BROM, IDOEIITELT HMO TO
TEIT HMO. LONER UTEIIIL: SHILE; GRIT TO CUT 8ROM. NOOEIITELT HIRO.
39.6-48.5 a - BUFFHO SANDSTONE INO SILTSTGNE: GIIIT TO GIIT Him. TERT FINE TO FIKE CIIINEO, NICICEOUS.
-ISO (130-155') IR6ILLICEOUS. MOOEIIITEIT HUD TO HIRO.
48.5-59.) 0 - SHALE: GRIT TO GIIT BRMN. IOOERITELT HMO. LOCILLT VERT HIID.
(159-194')
-JJO 59.1-65.2 n - BRUSH CREEK MARINE SHALE: DIRK GRIT. NOOERITELT HIRD: HUH OREINIC IITERIU [WON.
(194-214')
65.2-66.5 i (214-215' ) - BRUSH CREEK COAL.
66. 5-75. S « - CORINTH SANDSTONE: 6»IT TO TERT LIGHT MOM. TERT fINE GUIINED. TERT HMO; CONTIIHS 1 THIN LINE-
(215-249') STOKE IEO l« UPPER HILF OF UNIT.
-250
75.9-89.0 • - SHALE; GIIT TO GRIT SHOWN. SHOT. NOOEIITELT HMD; IKTERIEDS OF SILTSTONE.
(249-292' )
-300 89.0-95.7 l» - MilOJUWLiANDSlfiNi ; LIGHT SIIT TO LIGHT BUOIN. TERT FINE 5RIINED. OUMTOSE. TERT HIIO: EROSIONIL
(292-314' ) LOIER CONTICT.
95 7-100 9 « - UPPER FREEPORT LIMESTONE; CHIT LIMESTONE OTEIUIN IT CUT TO ERIT IRDIN SHILE.
(314-331' >
100 5-116 7 n, - INTERBEOOEO SHALE, SILTSTONE AND SANDSTONE; GRIT TO GIIT IR01N IND CIEENISH GRIT. TEIT
(331-383' ) FINE TO FINE SRIINED, IRGILLICEOUS. NICICEOUS. IQOERITCLT HMO TO TERT HMD; INCIEISE II FLOI
IT III.) . (3«3'>. TOTH FLOI 0.2 Is (3 IP).
116 1-117 3 in 1383-385' 1- LOWER FREEPORT "0" COAL.
117.3-128.9 n - INTCRBEODEO SHALE. SILTSTONE AND SIHOSTOSE: 8IIT TO GIIT IIOIN INO GREENISH GRIT, YERT
-400 (385-423' ) FINE GRIIKED, IRGILLHEOUS. NOOERITELT HIRO TO TEIT HMO; POSSIBLE THIN COIL SEIN NEII TOP
OF UNIT.
118.3-129.4 m (423-424. 5M - UPPER KITTANNING "C PRIME" COAL.
129 4-140 B IS - UPPER IORTHINGTON SANDSTONE: GNU TO GIIT lion. TEIT FINE TO FINE CHINED. OUIRTOH. TERT
4SO (424. 5-462') HIRD.
140.9-151.6 n - SHALE; GRIT 10 GRIT IROIN. MMEIITELT HIIO.
(462-497. &' )
151.6-IS2.2 m (497. 5-499. 51) - IIIOOLE KIITINNING "C" COA1.
3UU 152.2-153.8 n - LMER IDRTHINGTON SANDSTONE: HIT TO HIT IIOM. TEIT FIKE SIIINED. OUIRTJSE. TEIT HMO.
004') (X99.5-504.51)
Figure 17. Drilling log: Dewatering well P-2.
52
-------
SYSTEM
GROUP
FORHAT I ON
•ELL
CONSTRUCTION
0
DEPTH
•ETERS FEET
STEEL
CISIH
!S.(ci
'i.1
STEE
'.I1.1 Ill
15. ZM
I.D
(8"
HOOK-
flLL
PIC IE R
125-
150
IS1.0 •
(SO!1 )
DESCRIPTION
ELEVATION 584m (1916.3')
0-10.7 • - SiiJJBilBCJANDSHIHE.: LIGHT II01I TO SLIGHTLY REDDISH BROfl. OUIRIOSE, IOOERIIELT HIRO 11 VEUT HIRO:
(0-351 ) IEITHEIED IEM SURFACE. SOIE LOCH IODERATEH HEirHEBEO, OTHERf ISE FKESH; FEI IITERIFOS »f SHHE IRO
SHILEI SIIOSTOIE: EHOSIOIAL IOIER COITICT; IITER EICOUITERED IT 4.6 • (15').
10.7-24.4 HI - SHALE, LIHESTONE: BMT LIIESTOIE 110 LIKI SHILE 1DUIOED BI UPPER 110 LOIER SHILE UIIIS. UPPER
(35-60') SHILE: HOOEIITE lk.il 110 Mil. IOOERITEU HIIO: IITERIEOS OF LIGHT IROII (CRT FIIE ER1IIED
SIIOSTOIE. LO»EI SKILE: Gill 110 RED 11 Oil TO IIROOI (KTEISDUE RED «EO?); 11101 GRIT CUT
ISSOCIIIEO: IICREISE II FL01 IT II.I I (5I1 ).
25-T" ~\ 24.4-29.9 HI - IMtkO. 5>HOSTG)IE-SILTS70NE: Gill, HIT FIIE TO FIHE ORIIHED, IRGHUCEOUS, IOOEUTELT HIIO TO
(80-381) YERT HIIO.
-100
HOLE DIUETEI
21.8oKI-l.-5") 29 j^.j „ - SHALE; mi TO OIR< Mil. IOOERITELT HARD; SME IITERBEDOED SILTSTOIE IM SIIDSTOIE; IICREISE II
(9B-1431) FIOI IT 36 • (III1). IOTU FLOI 0.3 I 'i (5 BO*).
43 B-49.4 n - BRUSH CREEK HARINE SHALE: OAK GRIT, IOOEIITELT HIRO; ILICI OUIIIC NITEIIIL COIHOI.
(143-162')
50 ,. n . 48 4-49 7 m M67-I63M - BRUSH CREEK COAL
53.3 • 49.7-60.0 n - CORINTH SANQSHNi; GRIT TO U6IT Mil, IERT FIIE TB FIK GRIIIED. HIIO TO »ERT BIRO: I1HDR
(I7S1) (I63-I971) SHILE IT UPPER COITICT.
-200
60 0-72.2 n - SHAUzJUTSJISf,; SIIT TO SOT IROII IM) UtHT TO IODE«IT£ till. KOOERITELl HIRD. FISSIE;
(197-237') IITEIIEOS OF ERT LIGHT GUI TO LIGHT IROII. IERT flit TO FIIE BRIIIEO. OUHTdSE, IERT
(237-263') HARD.
80.2-85.6 m - UPPER FREEfORT UHSTONi: IITEIIIEOS OF LIIKSTOIE 110 LUEI SHAIE. 8ROIH TO GRII IDOfl. IOOER-
(263-28)' ) ITELT HIRO TO HIRO.
85.6-88.1 m - CIA1; M»l.
(28I-2891)
300 08.1-100.3 m - INTJgBfpDin SHALE. SILTSTONE ANO SANOSIONE: Mil TO GRIT IIOII. VEII FIIE GUI 1KB.
(2(9-329.') H6ILUCEMS, lOOEIAIEll HIID TO lEIf HAM.
IDO-taHHl 100.3-101.2 in (329-332') - LOIER FflEEPQRT "D" COAJ..
101.2-112.8 n - ljlTE»aipDEj_SH»LE. SILTSTDNE AND SANDSTONE: MIT TO ERIT IIOII. »£«! FIIE 10 FIIE 6RIIIEO.
-350 (332-370') OUIIUSE, IOOEII1ELT HIIO TO Hl«0; 1RDII TO CUT IIOII CUI I! TOP OF UIIT; IICIEASE II FLOI II
HOLE OIAKTEI ""'" " '"'''• ""'L fl0i °'!MM ' ! <8""' '
JH, [^J 112.8-113.4 • (370-372') - UPPEB HITIANHIN6 "C PRIM" COAl.
~*W 113.4-134.4 «1 - MgPjR 10KTHINCTON SANDSTONE; Mil TO till MOII. YERT FIIE TO FIIE GIIIIED. OUIITOSE.
(372-441') AICILLACEOUS. IOOERITELY HIID TO IERI HIIO; IITEIIEOS OF SILTSTOIE I ID SHILE.
-450
134.4-135.0 n (MI-4431) - jlOOLt KITUNIN8 "C" COAL.
135 0-150.0 • - Lj«LlMHlHSI8NJ«fi5TS!!I: "«. K«I HIE BRIIIIED. OUIRTOSE. KERT HIIO: HIM DIM GRIT
(443-492') * iiilLE IT «PPE» 110 LHER COITICTS; IICREISE II FIOI IT Ml.4 i (4101).
150.0-151.5 • (492-497') - ICIER KITTANNIMC "8" COAL
151 5-153.0 m - SflAli; Mil ID Mil IIOM.
(497-502')
Figure 18. Drilling log: Dewatering well P-3.
53
-------
— GROUP
f FORMAT I ON
IELL METERS
CONSTRUCTION
i
j
•
• i
<-
CJ
-
GLENSHA1 FORKA1
i'.
a
.-.
.
••
i
CIS lit i
1.0.
SIEEL
cis IK
II-.1,..
I.D.'
HOOniLL
PICIII
Ifi.JC"
1.0.
t».7')
25-
HOLE DIICTER
50-
!:!.H *
(110')
75-
HOlf OIIKTEI
K.3c»(WyB")
100-
125-
il. 1 i
DEPTH
—
:=.>-
• '''.•I
MMl
MM!
MM*
EET DESCRIPTION
-0 ELEVATION 566.3m (1857 8')
0-6.7 in - coj.LU.viLm; slum CLUE! SILT. IODEIATE aim 10 TELLDI tioii.
(D-221 )
6.7-21.9 TONE; HIT. VEUT FIIE BIltlED. OUAIIOSE. »EIT HAIO: HIM CUT
(438-485*) SHALE IEII USE; IICRE1SE IN FLO! IT III. 2 • (
-------
STSTEH
GROUP
[
.
> .
.1
,
•• '
i
• !
'I
•
1
s
•'
i •;
*
II;
™
L.
•' '
J
HIRNWI
CONS
STEEL
CASING
JO 3ci
a
(1")
STEEL
CASING
10 2on
I.I
(4")
HOOIULL
PICKER
10 2cm
ID
14")
UN
• ELL HETERS
IRUCTION
n n
5.9.
(19.3')
25-
NOLE OIUNTHR
17 ICK6-1 4"l
50-
| IB.! , 75-
(250')
100-
HOLE DIUEIER
l4.1oi(5-S 8")
129-
!
150-
156. I •
(514')
m
P—
y-~-
—
^^^
_
i\
••]
IG
-100
-150
-200
-Z50
-300
-350
-400
-450
-500
DESCRIPTION
ELEVATION 566 0» (1856 5')
0-25.3 HI - S>LTS6URE SUKOSTOKE; LIEUT lEUM BRUIN Tt SRAT 8(0111 (III LIGHT Mil. ICRY FINE 10 KDIUI CRIIIED
(0-B31 ) NOOEIITELT HUD TO »ERT mil; IEIIHEIED NEII SURFACE. SOW LOCAL IOOERATELI NE1THERED OTHIRIISE
FRESH; iota IHTERIEDS or SMILE no CLAYIT sine IIIH CIRIONICEOUS «>I[«IH; EROSIONAI L»I« co«nci.
25.3-J9.6 IB - SHALE. LUESTOKE: LIGHT ILIIISH SRAT LIKS1ME SOUNDED IT UPPER 110 LOHI SHILE UNITS. UPPER
(83-130', "Mlf: IMi Bill. MDEUTELT MID. LNCI SHUE: CUT. NODERinLT HIID; INTfRtEOS OF GRII
SILTSlOUt; ENCOUNTERED NITER IT 31.8 • (!!<'). n 3 I s (.5 u»).
39.6-46.5 m - BUFf»LO 5HNOSTONE: CRIT. IERT FINE 10 FIIE CIIINID. UDEIIIEU HlRO 10 HlRO
(I30-I591)
48.5-61 .0 II - SHUE: SLIBHTLT 1LUISH GUT, FISSILE. SOU TD NIOERITELT HIHD
(159-200')
6I.D-68.0 m - aftU5i.QltiJLJARJNj_SHiki: OH« MIT; KHENITELI KURD; IL«C« OIG»IC IITERIH CONNOII.
(200-2231)
68.0-68.3 01 C223-224' ) - BRUSH CHEEK CtUL.
68.3-72. 8 n - SKAL£| GRIT TO DIRI CRIT. NODERIIILT HIRD.
(224-239')
72.8-18.B IB - CORINTH S«HOSTOMt: LIEUT GUT BI»f» TO LISHT BRIT. KEIT Flli TO FINE SRIINED BUIRTOSE »E«T
(239-258') HlRO.
78.6-93.6 » - SH>Lt. SILTSTGNt: SHUE 0»E«LTI«t SUTSTONE: GRIDITIONIl CONTACT; LIGHT GUI! TO BUT; NODEKITEIT
(258-3D71) HlRO TO HARD; RED SHUE HEIR NIDDLE OF UNIT («HO«I«G RED BED').
93.6-IDO.O n - N*HONI KG S»NDSTO>iE | LIGHT GRIT TO CRIT IIOIN. IERT FINE EIIIICI. QUIRTOSE. HIIO TO IER1 N1RO-
(307-328') CIIOtTIONIL UPPER COHTICT
100.0-109.4 m - 5HUi; SRIT Tl Dill GRIT 110 GRIT MOM. IOOERITELT HIIO TO HIIO; LENSES OH NOOOLES OF HIRDI
(328-359') NITERIU IN UPPER POKTION OF UNIT (LIKSIDIE'. SIDEHIIE7).
109.4-MB.B IN - iiBlilM.; tin 10 GRIT INDIl. fERT FINE 10 FINE GRAIIEO. IODERATELT HlRO TO TERT HARD: IITER-
(359-390' ) (EOS OF SILISTOIE.
lit.6-119 2 n (390-391') - LMiB fRttPORT "0" COIL
U9.2-I3I.7 a - sniu: SUIT IIOIN. SOFT TO NOOIHTELT HMO; INTERIEOS of IERT FINE TO FINE GIIINEO SIIBSIONE-
(391-432' ) CONTAINS 0.61 . (11") COAL SEII IN LONE! HILF; INCREASE II FLOI IT 127 I > (1111) TOTAL
FLOI I .1-1 6 I s (29-25 »•).
131.7-132.1 n (432-433.5') - UPPER KirTAMNIIIC "C PRInt" C01,t
132.1-143.6 n - llEEEjjDRTJUifiijllLSAIIIlSIJSI; BRIT. »ERT FINE TO FINE ERIINED. «RT HIIO- IIIOR SHIEET
(433.5-471') INTERIEOS.
143.S-153.3 » - Mill; M«t TO OIRI GRIT. NOOERITELT HUD TO HIIO: HINOR CIR10RICEOUS NITEIIIL tSSOCIIlEO
(471-503') OIRI CRIT SHILE; IIT[R>EOS OF SILTSTONE.
153.3-154.0 « (503-505') - HtULl mTTANNIHJ ."C^JCM
154.0-I5S.7 » - lOItR IOR7HINGION SINDS10NE; G«IT. IEIT FINE EIIINiO. OUIRTOSC. >ERT HlRO
(5B5-5I41 )
Figure 20. Drilling log: Observation well OB-1.
55
-------
• .1 .IK
• i.Hiiui1
-I IJHIAi I (IN
• m HE IEH:
LUNMHULIIOH
M
*
M
t-1
I
1
-
&
.
S
j
m
<•
,
-
"
i
'
situ
CAS IN!
Ill II M
1.0.
STEEL
C1SIHG
10. h«
1.0.
14")
KWBItl
PICK!
lll.7c«
I.D.
- u u
IH.'i' )
25-
HOLI OllttTER
ir.knlW 4")
50-
iS.I •
(MS1)
75-
HOU DIAHEIER
100
425-
150-
161.1.
(5501 )
^
•MM
•vm
—
HP«V>
•1
f£M
-0 ELEVATION i?7.4i» (1894.5' )
Di SCSIPI ION
0-15.2 m - SALISBURY SANDSTONE; HUNT IRMN ID TfUOf IIMN mo NGMI1IE linn. FINE 10 11 jiin UIIIIEII ink
(0-bO1) sm mill. juio. gumo". IDOIIITIU NINO 10 nm HUH, •tiiwitn Hill suiun. iwiiii ion
•ODIRIIEL) ItimtltO. OUtlUlt fltn.
5JJ
15.2-35.4 in - SHALE AND LI « SI Wti ; CRIT LIKSTOIE IOUIDEO IT UPPER 110 LOIER SHU! UNIIS. UPPER SIHLE:
(50-II61 ) MODISH IIOfN AND CUT; IHIEIIEOS OF SANDSTONE IIIH SOW CAIIOHlCfOUS HAltRllL. LOIJ.II SHALL :
GRIT IU DARI GRII IIIH SHINE RED IIOIN TO NilOON IEII USE (IETERSOILE (ED HElT). NBUtHlELT
HARD; IHTER1IOS OF (ERT HIE CRAIIIO SANDSTONE AND SILISTONE; ENCOUHTEREO IATCI II 'I I m
100 ('»<'>. ' 1 I s (J if* I
35.4-3J.a ni - BUFFALO SANDSTONE; GRIT 10 SREENISH GRII. HIT FINE TO FIHE CRAIHED, IICICEOUS, IIGILLACEOUS.
HUB.
31.9-53.B n - SHALE; Gill 10 OIRI IRK. I01ERHELT K1RO; !OK IITEIIEDDED SILTSIOHS.
ISO HAHINi SHALE; Hill HIIT, HOOEIIIEIT HIRl, ILICk Mimic IITIIIIL -IMW
(176-204')
62.2-71.6 IB - COHIKIH SANDSTONE; GRIT TO LIENI CUT IIMK. rE«T HUE TO I III GRIINtO (lilt ml:, lu mill
(204-235')
71.6-85.3 m - SHALE, 1ILISTONI; SHHE IIEILIIM SILISTOKE. EIIDITIOIIL COKIACI; CHIT 10 CHIT MOM. KOOiK-
(235-280' ) ITELT HIM TO «IT HAIO; RED SHILE REIR IIOOLE OF UNIT («»HO»I»S RED lEOV).
85.3-91.4 in - MHONINL SANOSIOHE; IEII LIGHT GUT 10 LI«HT IRDIR, IERI FIHE ID FINE CHAIRED. rlHI KIIO;
3QO (280-300' ) " INCREASE III FUJI AT 10.S • (ID').IOIIL FLM 0.3 I i U 1(11)
91.4-112.8 B - IJtTERBEODED SHALE, SILTSIONE AMI SAHOSIPJtf; HAT Tl GUT 810m DIRK GRIT inn GREENISH
(300-370') GRIT. FIDE GIIKED. JRCILLACfOUS. HIID.
112.8-113.4 » (370-372') - LO«H FHiPORT "P" CCJAi.
113.4-125.0 n - INTERBEDOED SHHE. SIL1STONE AND SANOSIONE; CHIT TO GRIT 1ROIN. MflNlsn CKH mo lull
(372-410') CUT. «EI1 FIDE TO FIRE GIIINFD. IRCILLICEOUS. IOOERI1ELT HIM 10 (Elf H1RD: IICRIIU IN
FLW IT 18.3 » i iff i IOIIL ILO» II < I s (1 ,c»
125.0-125.3 m (4IO-4II1) - UfftR KITTAHNING "C PBIHC" COAL.
125.3-136.8 fll - UPPER IORTHING10N SANDSTONE: «» 'IRE TO FINE Sill IEO. OUIRIOSE. (IR( NARO.
(4I1-44D1)
136.9-147.2 m - SHAH; OIK G«IT TO GtU IROIN. SO* (ERT FINE GRAIIED SANDSTONE INTER1EOS.
(449-483' )
147.2-147.5 in (4B3-4B4.51) - JIDDLE KITTAKHIHG "C" COAL.
147.5-165.5 in - LOIER 1URTHINGTON SANDSTONE; (NO HE THIN OF CUTIIDGS). (EIT IIIRO, LOCILLT NOOERITELT
(414.5-543') HAIE: SOFTER SHILET IAIERIAL II TOP INO RISE OF UNITS.
165.5-167.0 n (543-548') - LOKER »|IIANINt "6" COAL; (NO RETURN OF CUTTIUGS).
167.0-167.6 in - SHALE; (NO RETURN OF CUTMNCS).
(548-550')
Figure 21. Drilling log: Observation well OB-2.
56
-------
3£
Tf.
i*j
.
r
i .
.
i '
. •:
I
| GLENSHAI FORMTIQN
ROUP
lT'°,kEU DEPTH
CONSTRUCTION DETERS
STEEL .
CAS INC
20.301
1.0.
. 5.0 •
(IK1)
25-
HDlf OlAKTtl
50-
'£'l'|i
••A^M
FEET DESCRIPTION
-0 ELEVATION 578.1m (1896.6')
0-H.6 m - SAylSeilRC SANDSTONE: LIGHI GRIT 10 LIGHT BRIT BIOH, »ERT FIM£ TO FIKE CRAKEO. OAIARTOSE. IOOERITELI
(0-6D »A«0 TO IER* HARD; IRTERBEDS OF SHILE.
|g 5-36 6m- SHALE SANDSTONE AND LIMESTONE: GUT TO BLUISH GRAI LIWSIOIE AND LIIEI SHALE J«0 CLAT UK9ERLAIR
(81-120') BY SHALE IITH SOtt SIKlsriKt; GRIT TO LIGHT BRO
-------
SECTION 7
PILOT DEWATERING PROGRAM
METHOD OF APPROACH
The objectives of the dewatering program were to intercept a major portion
of the groundwater inflow to the Main G study area and thereafter to analyze
the cost-effectiveness of mine dewatering as a means of controlling AMD. The
approach used was to construct vertical wells adjacent to the mine openings
to intercept groundwater before it could be degraded in the mine. The wells
were located along a line parallel to the mine face in an attempt to create a
hydraulic drawdown barrier against groundwater movement toward the underground
study area. Despite the complexities of the groundwater flow system (described
in Section 6), the results of the pilot program demonstrated that well dewater-
ing can be effective and also pointed out some of the practical problems in-
volved. The program further demonstrated that the untreated water pumped from
a dewatering system can be of significantly higher quality than treated mine
water and can be discharged directly into a receiving stream.
Pumped wells were used for this study because they provide more information,
and create less disruption of mining and less safety hazard than other possible
methods. A discussion of relevant alternative dewatering methods is presented
in Section 8.
DEVELOPMENT
The program, which was used initially to develop baseline data, as well as
to prepare for the dewatering program, was carried out in three phases. The
initial phase consisted of the development of an underground monitoring facil-
ity. In the second phase surface geologic mapping was performed and one test
or pumping well and two observation wells were constructed. The purpose of
this phase was to determine the feasibility of a dewatering scheme using wells,
and to provide the necessary data for locating additional test and observation
wells. This phase was completed by the construction and pump testing of these
additional wells for the pilot dewatering program. The third and final phase
was the selection of appropriate pumps and generator. These activities were
detailed in the preceding section.
OPERATION
Based on data obtained during the preparation phases of this study, the
pilot well dewatering program was scheduled to be operated for two or three
weeks. During operation, monitoring and sampling on the surface and in the
mine were carried out on a continuous basis. Equipment for both of these
58
-------
operations was the same as that used during the pump testing. Drainage from
the three dewatering wells was discharged into a small basin located near
P-3 and allowed to flow by gravity through a small culvert under the railroad
embankment and down to the nearby stream. To prevent possible seepage losses
into the subsurface, the basin adjacent to the culvert was plastic lined.
The pilot well dewatering program began with continuous pumping on July
13, 1977. Total pumpage for all three wells initially was only 2.87 t/s
(45.5 gpm); Well P-2 was yielding only about 0.2 tjs (3 gpm) and the pumps
on P-l and P-3 were producing only about half their capacity. The effect on
mine inflow was only slightly greater than that seen in the individual pump
testing of the wells. Pumps were shut off on July 16, and the wells were
allowed to recover. Well P-2 was abandoned as a pumping well and the 3.7 kW
(5 hp) pump in P-2 was pulled and reinstalled in P-4, which had been used as
an observation well (OB-4) up to that point. Pumping resumed on July 18, and
the total pumpage was increased to 4.6 £/s (73.5 gpm). Pumping continued at
this rate until July 20, when once again the pumps were shut off. This time
the cause of shutdown was intense rainfall during the nights of July 19 and
20, accompanied by widespread surface flooding. As a consequence of this
storm the underground workings of the Lancashire No. 20 mine were completely
inundated. Pumping at the wells and all other field operations was tempor-
arily shut do*wn until the mine could be pumped out. Total rainfall in the
area from the storm was greater than 30 cm (12 in.) in a 24-hour period.
Water levels in wells started rising before the mine was flooded, indicating
recharge from direct surface infiltration of rainfall.
Field operations resumed in late August and the pilot well dewatering
program was again started on September 16. Prior to the start of pumping,
wells were cleaned to flush out any fine-grained sediments which might have
flowed into the wells during the storm. In the mine the underground transport
of water was also reorganized following the clean-up of the mine, and monitor-
ing equipment underground was relocated to compensate for this change. With
the continuation of pumping in mid-September, minor shake-down problems
occurred during the first couple of days of pumping. Pumps were shut down
twice because of minor problems with the generator and pump control box. After
the problems were corrected, pumping was continuous from September 23 until
October 7, 1977.
MONITORING
Prior to the start of dewatering, procedures were determined for monitoring
of well discharges and mine inflows and for sampling of both surface and under-
ground waters for water quality monitoring. Equipment and methods used for
these operations were similar to those employed during the pump testing and
baseline data collection periods.
Flow Measurements
Flow measurements were made on a continuous basis for the inflows to the
Main G study area and for discharges of dewatering wells, along with water
level measurements in observation wells and pumped wells. Stevens continuous
float-operated water-level recorders were installed in the observation wells
59
-------
to monitor changes in water level. Water levels in pumping wells were mea-
sured on a regular basis with an electric water level indicator similar to
that used during pump tests. Pumpage rates in the dewatering wells were re-
corded hourly, initially using a 5 cm (2 in.) digital flow meter. These were
removed 3 days after the start of pumping because of jamming from silt and
sand sized rock particles. A 20 i (5 gal) bucket, and later a 210 t (55 gal)
drum, were used to measure the discharge from each well. This method of flow
measurement was considered to be adequate because the discharges were rel-
atively constant with only gradual changes, and the discharge from each well
was only 1.3 to 2.5 £/s (20 to 40 gpm). Errors in measurement probably were
less than 10 percent, which is adequate for this study.
The underground monitoring system described earlier was utilized to obtain
mine inflow data for the section of the mine being studied. Water meter
readings were taken once each shift (three times per day) by supervisory per-
sonnel working in the Main G area. Meter readings were recorded each shift
and then transferred to a master data sheet in the foreman's office each day.
These flow measurements were collected daily by Wahler or ENCOTEC personnel.
Occasionally personnel from the study team would go underground and take
measurements to check the readings being taken by the mine operators.
The Stevens recorder was capable of eight days' unattended data recording.
Each week a member of the study team went underground to pick up the recorded
data and reset the Stevens recorder. Other equipment was also checked at this
time along with a general inspection of the test area of the mine. As des-
cribed previously, the monitoring system in the Main G study area had to be
rearranged after the July 19-20 storm. However, the same monitoring schedule
was used when dewatering operations were resumed in September.
Water Quality Monitoring
During the pilot dewatering operation, water samples for analysis were
collected from the Main G study area, the dewatering wells and from two
stations on Laurel Lick Run, above and below the discharge of the wells, as
shown on Figure 23. Grab samples of the creek water were collected 7 times
during the September dewatering period spaced at 2 to 4 day intervals.
Grab samples were collected from each pumping well throughout both de-
watering periods in July and September. Analyses of these samples provided
data to detect changes in water quality in any of the wells during the de-
watering periods. Also, these data indicated the acceptability of the well
water for direct discharge to ,the creek according to the discharge require-
ments of the Commonwealth of Pennsylvania.
Water in the Main G study area was sampled in July from the discharge
pipe from the L-2 sump, and in September the sampling station was moved to
sample discharge from the L-2 pool as described in Section 6. Water samples
at both locations were obtained using the Protech composite sampler. The
:sampler was set to collect a-fixed-volume sample each hour and a 24-hour
composite sample was generated. Each day during the dewatering period
(including the recovery period), mine personnel would change sample contain-
ers and bring the sample to the surface. Occasional grab samples were
60
-------
$00 METERS
100 FEET
Figure 23. Location map:
CONTOUR INTERVAL 20 FEET
Surface water sampling stations,
61
-------
collected to check the results of the composite samples. In addition, grab
samples of the F-14 pipeline to the treatment plant were obtained during both
dewatering periods. These samples provided analyses to compare with baseline
data.
RESULTS
-f
Effect of Pumping on Mine Inflows
The first pilot dewatering started on July 13, 1977, pumping from Wells
P-l, P-2, and P-3. The yield of the wells was increased gradually to a max-
imum total of 2.9 t/s (45.5 gptn). This yield was disappointing because P-l
and P-3 were both expected to yield 2.5 to 3.2 £/s (40 to 50 gpm) each, based
on previous pump tests. Well P-2 was not expected to produce much water, as it
did not, and the pump was later removed. Pumping continued for about 3 days
and was stopped on July 16. During this period flows in the Main G study
area of the mine had decreased from 6.7 t/s (107 gpm) to 6.0 t/s (96 gpm), or
10 percent, as shown on Figure 24. About 24 percent of the water pumped from
wells was intercepted or diverted from the mine. These results were improved
(Figure 25) by changing equipment.
The 3.7 kW (5 hp) pump was removed from P-2 and installed in OB-4, there-
after designated P-4. Pumping resumed on July 18, and total pumpage was in-
creased to 4.64 t/s (73.5 gpm). On July 19, flow in the Main G study area,
which had increased to 7.0 t/s (111 gpm) just prior to starting the pumps,
rapidly decreased 28 percent to about 5.1 t/s (80 gpm). This indicated that
42 percent of the water pumped from the wells was diverted from mine inflows.
The operation was interrupted by mine flooding, but these results were encour-
aging because the percentage of water intercepted should increase with time.
Dewatering was again resumed on September 16, 1977 after cleaning the
wells and resetting pumps. Minor problems with the generator caused two
breaks in pumping between September 16 and September 23. After this time the
wells were pumped continuously until October 7, 1977. The average total
pumpage was approximately 5.2 t/s (82 gpm), varying from about 6.0 t/s
(95 gpm) near the start and decreasing to about 4.7 t/s (75 gpm) at the end
of the pumping period. The pumping rates for each well are shown on Figure
26; total surface pumpage is shown on Figure 27; and water levels on Figure
28. Water levels in observation wells showed a mixed response to pumping, as
shown on Figures 29 through 32. Well OB-1 responded as expected, with gen-
erally consistent declines during pumping except near the end of the period.
Well OB-2 declined rapidly at first and then more or less stabilized. Water
levels in P-2 declined slightly and then rose. Well OB-3 had essentially no
response to pumping.
Inflows to the Main G study area (Figure 25) decreased rapidly at first
but showed a slight decreasing trend after September 25. Inflows ranged be-
tween 4.4 and 4.8 t/s (70 to 76 gpm) after this date and quickly recovered to
7.1 t/s (112 gpm) after pumping stopped. The average inflow rate from
September 25 through October 6 was 4.62 t/s (73.2 gpm) with a standard
deviation of ±0.12 t/s (±1.92 gpm). The post-pumping monitoring period, after
inflows had recovered, from October 8 through October 22, showed an average
62
-------
7.0
6.0
5.0
4.0
START P-1-
-START P-3 START P-3 &
START P-1 P-4 (OB-4),
• START P-2
SHUT OFF P-1. P-2. & P-3
PUMPING DISCONTINUED
(MINE FLOODING)-
I I I
I I L I I
110
100
90
80
70
9 10 11 12 13 14 15 16 17 IB 19 20
July
TIME (Days)
Figure 24. Average daily mine inflow - July 1977 dewatering.
63
-------
7.0
5.0
4.0
STANDARD DEVIATION
PUMPING STARTED
SEPT.16 AT 9:30 A.M.
PUMPS P-1, P-3 AND P-4 SHUT OFF
SEPT. 18/19. 8:45PM-5:OOPM
PUMPS P-1 AND P-3 SHUT OFF
SEPT. 22/23. 3:OOPM-11:OOAM
PUMPING ENDED
9:00 A.M. OCT. 7
V/—V/STANDARD DEVIATION
I , I
1
1
I , I
1
1
I
I
I
HO
100
80
70
16 IB 20 22 24
SEPTEMBER
26 26 30
8 8
10 12 14 16 18 20 22 24
OCTOBER
TIME (Days)
Figure 25. Average daily mine inflow - September 1977 dewatering.
-------
2.5
2.0
1.0
• HEAVY RAINFALL ON NIGHT
OF SEPTEMBER 25/26
PUMPS P-1. P-3 AND P-4 SHUT-OFF
SEPT. 18/19. 8:45 P.M.-5:00 P.M.
+ •* + * PUMPS P-1 AND P-3 SHUT-OFF
SEPT. 22/23. 3:00 P.M.-11:00 A.M.
40
35
25 -5
C9
20
17 t9 21 23 25 2? 29
September
TIME (Days)
1
3 5
October
Figure 26. Average daily pumping rates - September 1977 dewatering.
6.0
5.0
4.0
~ 3.0
2.0
1.0
75
50
25
17 19 21 23 25 27 29 1 35 7
September October
TIME (Days)
Figure 27. Combined daily pumpage rate - September 1977 dewatering.
65
-------
550
500
450
400
PUMPING STARTED
SEPT. 16 AT 8:30AM
PUMPING ENDED
8:00AM OCT. 7
• + PUMPS SHUT OFF
I . I . I . I . I I I I I I I < I ' 1 L_J l_J L_J I I I I > J
1800
1700
1600
1500
1400
7 9 11 13 15 17 19 21 23 25 27 23 1
SEPTEMBER
TIME (Days)
3579
OCTOBER
Figure 28. Water level elevations in dewatering wells - September 1977 dewatering period,
-------
- 2.0
4.0
6
r—^ PUMPING STARTED
lSEPT. 16 AT 9:30AM
PUMPS P-1. P-3 AND P-4 SHUT OFF
SEPT. 18/19. 8:45PM-5:OOAM
PUMPS P-1 AND P-3 SHUT OFF
SEPT.22/23, 3:OOPM-11:OOAM
HEAVY1 RAINS ON NI6HT
OF SEPT. 25/26
PUMPING ENDED
9:00AM OCT. 7
llllllllllllllllllll.il
17 19 21 23 25 27 29 1 3 5 7
October
TIME (Days)
10 2
15
September
Figure 29. Drawdown: Observation well OB-1 - September 1977 dewatering,
0
2.0
PUMPS P-1 AND P-3 SHUT OFF
10
4.0
6.0
8.0
PUMPING ENDED
9:OOAM OCT. 7
I 1 I
PUMPS P-1. P-3 AND P-4 SHUT OFF
i 1 i 1 i I i I i I i I
15
20
25
9
Figure 30.
17 19 21 23 25 27 29 1 3 5 7
September October
TIME (Days)
Drawdown: Observation well OB-2 - September 1977 dewatering,
67
-------
2 -
oc
0
HEAVY RAINS ON NIGHT -%j\
OF SEPT. 25/26 N \
PUMPIN6 STARTED \
SEPT. IS AT 9:30AH V^-,
V~— , , , . T 1
I 1 t 1 1 1 1 1 , 1 1 1 1 1 1
17 19 21 23 25 27 29
SEPTEMBER
PUMPING ENDED -i
9: 00 AM OCT. 7
.s • • •-~4__--
i i i i i i i i
13579
OCTOBER
- 5
0 -
- 0
TIME (Days)
Figure 31. Drawdown: Observation well OB-3 - September 1977 dewatering.
-14.0
-12.0
-10.0
" -8.0
-B.O
-4.0
-2.0
2.0
HEAVY RAINS ON NIGHT
OF SEPT. 25/26
PUMPING ENDED
9:OOAM OCT. 7
PUMPING STARTED
SEPT. 16 AT 9:3QAM
I i ! i t i I i 1 I I i I i I i I
17 19 21 23 25 27 29 1 3 5
SEPTEMBER OCTOBER
TIME (Days)
-40
-30
•20
-10
7 9
Figure 32. Drawdown: Observation well P-2 - September 1977 dewatering.
68
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flow of 6.9 £/s (110 gpm) with a standard deviation of ±0.25 Us (±4 gpm) .
These values are quite comparable with all of the other background data, which
showed an average inflow rate of 7.1 £/s (112 gpm), ±0.59 t/s (*9.4 gpm).
As can be seen in Figures 26 through 32, all the pumping and observation
wells, with the exception of OB-2, show the effects of recharge from direct
precipitation. The pumping and inflow data indicate that the mine study area
inflow was decreased by about 34 percent based on average flow rates. It may
have been as much as 38 percent, based on measurements at the end of the
pumping period and the first inflow measurements after full recovery. These
data indicated that 45 percent of the water pumped from wells comprised inter-
cepted or diverted mine inflow, based on average rates for the pumping period.
Theoretically, this percentage should increase with time, and at the end of
the test up to 56 percent of the water pumped was diverted from mine inflows.
Projection of these data on the graph developed by Theis and Conover (1963)10
to 120 days of pumping indicates that up to 80 percent of the water pumped
would be diverted from mine inflows. Recharge events from storms or recharge
from streams would reduce these percentages. Therefore for purposes of es-
timating full scale mine dewatering, it was assumed that 50 to 80 percent of
the water yield from dewatering wells would be diverted from mine inflows.
During the dewatering operation, well yields were still disappointing.
Only Well P-3 was close to the maximum available drawdown, and that could not
be sustained during the latter part of the pumping period. Well P-l had more
than 30 m (100 ft) of available additional drawdown. The 5.6 kW (7.5 hp)
pump in each of these two wells was not able to produce at its rated capacity,
which was in excess of 2.5 l/s (40 gpm) with a pump lift of 155 m (510 ft).
In contrast, the 3.7 kW (5 hp) pump performed relatively well throughout most
of the dewatering program and during earlier pump tests. However, the yield
of P-4 gradually decreased along with a small increase in drawdown. The im-
pellers on the 5.6 kW (7.5 hp) pumps may have been worn, reducing their ca-
pacity, and/or gas blocks may have developed to retard well yield. All three
pumping wells produced methane gas, which was detected qualitatively but the
volume was not measured. The pumped water obviously had a large amount of gas
entrained in the discharge.
Water Quality
A summary of various water quality parameters measured during the two de-
watering periods (July and September) is given in Table 5, and water quality
data for the two pumping periods are shown in Tables 6, 7, and 8, located at
the end of this section.
Water quality was monitored in each of the three pump dewatering wells
throughout both dewatering periods. There were some differences between the
various wells. The water in Well P-l was lower in pH, alkalinity, and dis-
solved solids than the water in the other two wells, while Well P-4 contained
water with higher levels of sulfate than the other two wells. Well P-l water
showed a noticeable decrease in calcium, magnesium, and iron levels during
the September 1977 pump dewatering period. Calcium and magnesium levels also
decreased in Well P-3 water during this period. Total iron, however, was
more variable in this well (P-3) and did not exhibit the general downward
69
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trend, although ferrous iron values did decrease. The water from Well P-4
did not show this decreasing trend in the level of certain chemical components
and was generally consistent over both the dewatering periods.
Water quality varied somewhat from well to well. The water in Well P-l
was lower in pH, alkalinity, and total dissolved solids when compared to the
water in Wells P-3 and P-4. Well P-4 water had the highest pH and dissolved
solids along with almost twice the level of sulfate (i.e. 41 mg/£ compared
to 22 to 24 mg/£). Iron levels in the well water at P-4 were the lowest of
the three wells and it was the most similar in character to mine inflows in
the underground study area. The variability in water quality between these
three wells was probably due to the relative amounts of water produced from
the deeper parts of the wells. P-4 probably produced a higher proportion of
its water near the "B" coal seam from fractures and from the seam itself.
Also, wells with better hydraulic connection with the surface (through frac-
tures) would tend to yield less mineralized water.
The quality of the water from all three dewatering wells was acceptable
for direct discharge to a surface stream during the dewatering operation.
Iron concentrations during the periods of dewatering were less than 1 mg/£,
except during the initial day of pumping. The somewhat elevated iron values
during the first day of pumping may have been due to contamination from the
casing and other materials in the wells. After the first day, iron concen-
trations in each of the dewatering wells remained low, although they were
somewhat variable. The pH of the well water was acceptable for direct dis-
charge and did not vary significantly during the study. The good quality of
well water was an important aspect of the dewatering program. Direct dis-
charge of the intercepted groundwater to the receiving stream is a major
benefit of any dewatering scheme. The elimination or reduction in the amount
of drainage water to be treated is highly desirable from environmental, oper-
ational, and economic perspectives. Water pumped from the dewatering wells
was discharged directly to Laurel Lick Run during the dewatering program and
caused no significant changes in the water quality of the creek. Laurel Lick
Run, which received the discharges from the dewatering wells, was weakly buf-
fered with an average pH of 6.8 during the dewatering operation. Hence the
creek water was highly susceptible to fluctuating pH values due to this low
alkalinity. On one occasion during the background studies, the pH of the
water in the creek dropped to 3.2, indicating the probable release of some
acid-containing materials somewhere upstream. Table 7 shows water quality
data for Laurel Lick Run at Stations A and B, above and below the dewatering
wells.
A comparison of the overall quality of the creek with that of the well
waters being discharged during the dewatering program showed that in general
the well waters were more stable (i.e. have higher alkalinity) and were lower
in sulfate (based on a composite concentration level from the three dewatering
wells) than the creek water. Iron values were relatively similar in the well
waters and creek, while total dissolved solids were somewhat higher in the well
water discharges. With the possible exception of dissolved solids, the well
water discharges were of comparable or better quality than the creek water.
70
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The discharge of the well waters would be of some benefit to the creek in
terms of pH and alkalinity, as these discharges would add buffering capacity
to the creek.
A full scale dewatering program at this mine would produce sufficient
quantity of water from the dewatering wells to influence the creek. However,
the influence on the quality of the creek water would be positive, since the
well water would provide a buffering effect (as described above) and iron
concentrations would remain basically the same. The only potentially adverse
effect of the dewatering well discharges on the creek water would be a slight
increase in dissolved solids. The dissolved solids levels in the water from
the dewatering wells range from 124 mg/£ to 178 mg/£. These levels are not
excessively high and would not cause major water quality problems in the
creek.
There are possible uses for water intercepted by dewatering wells other
than to discharge it to a surface receiving stream. Assuming good water
quality, the intercepted groundwater could possibly be utilized by the mine
itself or by nearby farms and communities. These other types of usages
(irrigation, water supply, etc.) could provide a direct secondary benefit to
those living near mining operations.
During both the July and September dewatering periods, water samples were
collected from the L-2 monitoring station. In addition, samples of the roof
waters at L-4 were collected during the July period to see if there was any
deterioration between the two sampling points. Comparison of the data on
samples collected on the same days in July (see Table 8) indicated there was
some degradation in sulfates, iron and total dissolved solids. However, the
degradation was not substantial, and sampling at L-4was discontinued. As can
be seen in the summary of results (Table 5), iron and sulfate concentrations
showed the largest fluctuations throughout the study, with variations of *50
percent or more about the mean. Other parameters such as pH, alkalinity,
specific conductance, and total dissolved solids were relatively consistent.
The water quality of the inflow to the Main G study area was not affected
by the dewatering operation. A comparison of the background data with data
collected during the dewatering periods shows virtually no change in any of
the parameters monitored. Water quality in the underground study area was
good throughout the dewatering program and remained good after the dewatering
pumps were shut down.
During the dewatering operations in July and September, additional sam-
ples were collected from the F-14 discharge line (influent to the treatment
plant). Analyses of these samples are presented on Table 8. The pH and
alkalinity values for the F-14 line tended to be higher during the Fall of
1977 than the comparable baseline data collected in the Fall of 1976.
Sulfate concentrations did not appear to change appreciably between these
periods. Iron values fluctuated considerably from 2.7 mg/£ to 18 mg/£ during
the September/October 1977 survey but showed an overall decrease when com-
pared to 1976 values. Therefore, there was some improvement in quality be-
tween the background and dewatering periods. This improvement may have been
due to increased flows from the K section of the mine. A longwall panel
71
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encountered high inflows and this water was pumped directly to the F-14
sumps. This movement provided relatively little of the exposure needed for
degradation in quality.
The pilot dewatering program did not influence the quality of water
pumped to the treatment plant or have an impact on the treatment of the mine
water. This was because only a small percentage (5 percent) of the total mine
water discharge was affected by the dewatering study. A full-scale dewatering
operation would affect the mine water treatment plant in more substantial
ways. It would reduce the capacity requirements of the existing water treat-
ment plant and reduce tankage required for mixing and chemical storage as
well as the land area'required for settling ponds. The quality of the non-
intercepted mine drainage water, on an absolute concentration basis, would
probably become poorer as drainage from old workings, etc., became a more
significant portion of the total treatment plant flow. The total mass flux
of pollutants (acid, sulfate, iron, etc.) from the mine would be reduced.
This would be due to the decrease in the quantities of water contacting mined
areas and subsequently becoming contaminated. Thus, a reduction in chemical
usage at the water treatment facility could be expected. This reduction
would lessen the amount of chemical sludge produced and the associated prob-
lems of sludge disposal.
The benefits of decreased water treatment plant size, together with re-
ductions in chemical usage, extend beyond the obvious economic advantage.
Lower chemical usage requirements are advantageous both from an environmental
and from an energy viewpoint. Surface receiving streams in the area of de-
watering operations would be receiving a more natural and generally better
quality water (assuming good quality water from the dewatering wells). Some
energy savings would be realized by a reduction in both the production and
transportation of the chemicals required at the water treatment plant. Re-
quirements for sludge disposal would be lower, resulting in less land usage
and a lowering of the potential pollution hazards involved in handling sludge
and supernatant sludge.
72
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TABLE 5. SUMMARY RESULTS - WATER QUALITY
Location
WELL
WATERS
RECEIVING
STREAM
UNDERGROUND
STATIONS
Dewatering
Period
Baseline
Juiy Program
P-l
P-3
P-4
Sent. Program
P-l
P-3
P-4
Baseline
Station A
Sept. Program
Station A
Station B
Baseline
Roof
L-2
July Program
Roof
L-2
Sept. Program
L-2
JH
Av3
8.3
7.4
7.7
8.0
6.8
7.3
7.6
6.98
(6.35
6.84
6.77
8.2
8.0
8.0
8.0
7.8
«£
—
0.2
—
0.2
0.1
0.2
0-4 c
1.59)c
0.3
0.3
0.2
0.1
0.3
0.2
0.1
Alkalinity
CaCO,
Av
117
89
108
110
70
104
107
16
15
16
118
114
111
122
114
SD
^_-_
—
12
—
15
12
4
6
5
5
9
6
15
6
5
Specific
Conductance
us/cm
@25°C
Av
264
195
230
—
176
337
267
214
166
146
263
260
300
265
265
SD
--
16
—
37
21
19
120
51
78
46
11
9
9
14
Sulfate
wit
Av
19
19
18
39
22
24
41
100
43
45
31
50
28
18
32
SD
—
3
—
11
10
3
70
29
21
13
22
9
1
3
Total
Dissolved
Solids
Av
156
143
147
182
124
161
178
150
126
124
213
222
210
175
184
SD
—
3
—
23
7
7
26
41
44
21
10
11
3
7
Ferrous Total
Calcium Magnesium Iron Iron
mg/£ mg/£ mg/£ mg/£
Av SD Av SD Av SD Av
__ __ — __ -_ __ 0.007
_ „ __ __ o.09 — 0.47
_ _ _„. ^_^ __ | 37
— — — 0.02 — 0.10
__ __ _ — o.24 0.30 0.48
__ __ __ Q.12 0.18 0.81
__ _
-------
TABLE 6. PILOT PROGRAM: WATER QUALITY DATA-WELL WATERS
Date
Location 1977
Well P-l 7/15
7/19
9/10
9/17
9/18
9/20
9/22
9/24
9/27
9/30
10/4
Well P-3 7/12
7/13
7/15
7/19
9/16
9/17
9/18
9/20
9/22
9/24
9/27
9/30
10/4
Well P-4 7/19
9/16
9/17
9/18
9/20
9/22
9/24
9/27
9/30
10/4
pH
7.4
7.3
6.50
6,90
6.90
6.75
7.00
7.02
6.39
6.91
6.71
7.5
7.6
7.8
7.8
7.10
7.30
7,30
7.30
7.55
7.20
7.39
7.39
7.42
8.0
7.45
7.80
7.80
7.80
7.65
7.59
7.31
7.42
7.38
Alkalinity
CaCO,,
J — —
88
89
58
74
80
62
80
84
36
78
74
90
113
111
117
78
92
97
106
110
109
108
113
119
110
98
105
108
109
109
108
110
105
107
Acidity
CaCO,
J
—
9
14
—
—
8
13
3
10
11
—
—
—
—
7
7
—
—
7
9
10
8
7
-
8
4
—
—
3
6
8
8
8
Specific
Conductance
"25°C
195
—
190
210
210
140
190
200
98
190
160
212
242
235
—
210
240
240
200
240
230
250
260
260
-
24C
270
260
230
280
280
280
280
280
Sulfate
m/t
20
17
48
30
25
18
16
17
16
15
17
23
16
17
16
48
29
26
21
24
18
17
17
18
39
46
36
40
38
39
39
42
42
44
Total
Dissolved
Solids
m/l
152
133
140
141
152
104
125
135
78
125
114
146
146
152
145
157
167
166
149
163
159
155
168
169
182
191
183
184
171
178
174
175
178
169
Calcium
—
—
22
15
13
11
9
10
6.6
8.9
8.9
—
—
—
—
15
11
10
8.8
7.4
7.8
6.9
7.1
5.5
-
25
23
24
25
24
26
26
24
26
Magnesium
mg/i
—
—
5.7
3.9
3.6
3.0
2.6
2.7
2.2
2.4
2.5
—
—
—
—
3.9
2.9
2.7
2.4
2.1
2.2
2.1
2.1
1.9
—
6.1
5.8
5.8
5.6
6.1
6.0
6.0
6.0
6.1
Iron
0.10
0.07
0.85
0.55
0.42
0.17
0.08
<0.02
0.02
<0.02
<0.02
0.20
0.11
0.03
0.07
0.50
0.04
0.35
0.08
0.05
<0.02
<0.02
0.03
0.02
0.02
0.02
<0.02
<0.02
<0.02
0.04
<0.02
<0.02
<0.02
<0.02
Iron
0.58
0.35
1.2
0.88
0.73
0.54
0.24
0.18
0.25
0.09
0.18
3.54
0.88
0.58
0.49
1.1
0.76
0.95
0.83
0.77
0.65
0.77
0.52
0.92
0.10
0.71
0.14
0.14
0.13
0.14
0.13
0.12
0.39
0.12
Grab samples used in establishing parameters.
-------
TABLE 7. PILOT PROGRAM: WATER QUALITY DATA - RECEIVING STREAM
Location
Laurel Lick Run
Station A
Laurel Lick Run
Station B
Date
9/17
9/20
9/22
9/24
9/27
9/30
10/4
9/17
9/20
9/22
9/24
9/27
9/30
10/4
PH
7,35
7.00
6.90
6.85
6.51
6.57
6.70
7.30
6.90
6.90
6.74
6.40
6.62
6.57
Alkalinity
25
16
16
16
11
12
12
26
16
16
18
11
12
12
Acidity
CaCO,,
3
-
3
4
4
4
4
4
-
3
4
4
4
3
Specific
Conductance
PS/ cm
<§25°C
270
150
180
170
120
140
130
270
140
180
170
110
140
130
Sulfate
«*/£
95
44
53
48
22
34
35
88
43
51
46
23
33
34
Total
Dissolved
Solids
mult
206
119
127
147
86
105
94
214
135
132
120
84
96
90
Calcium
21
16
16
15
10
12
12
22
14
16
14
10
11
11
Magnesium
7.3
4.4
5.2
5.3
3.7
4.2
4.2
6.9
4.1
5.1
4.9
3.6
4.2
2.9
Ferrous
Iron
<0.02
<0.02
<0.02
0.04
0.02
<0.02
0.04
<0.02
<0.02
<0.02
0.04
<0.02
<0.02
0.03
Total
Iron
mg/X
0.42
0.63
0.44
0.42
0.35
0.32
0.34
0.34
0.70
0.40
0.31
0.38
0.25
0.21
Grab samples used in establishing parameters.
-------
TABLE 8. PILOT PROGRAM: WATER QUALITY DATA - MINE WATERS
cr>
Location Date
Main G Study Area 7/12
Roof Water 7/13
7/15
7/19
Underground 7/12
L-2 Station 7/13
7/15
7/19
9/15
9/16
9/17
9/18
9/19
9/20
9/22
9/24
9/27
9/30
10/4
10/8
10/10
Treatment Plant 7/13
Influent Line .
F-14 //1S
9/17
9/20
9/22
9/24
9/27
9/30
10/4
10/8
10/10
Sample
Type
Grab
Grab
Grab
Grab
Grab
Comp.
Cotnp.
Comp.
Grab
Comp.
Comp.
Coup.
Coup.
Comp.
Comp.
Comp.
Comp.
Comp.
Grab
Grab
Grab
Grab
Grab
Grab
Grab
Grab
Grab
Grab
Grab
Grab
Grab
Grab
Alkalinity
mg/£
pH CaCO,
7.9
7.7
8.1
8.1
8.1
7.6
7.9
7.4
7.50
7.75
7.85
7.80
7.80
7.80
7.85
7.82
7.82
7.70
7.78
7.81
7.82
6.8
7.6
7.00
6.80
6.80
6.85
6.71
6.61
6.75
6.69
6.79
j—
124
116
120
129
120
115
119
89
113
114
116
119
104
108
119
116
116
116
108
116
113
82
100
85
102
68
92
106
76
140
72
83
Acidity
ng/t
CaCO
j
—
—
—
—
4
—
—
—
5
5
5
5
6
5
9
—
—
7
—
33
35
27
29
23
27
22
Specific
Conductance
US/cm
-------
SECTION 8
MINE DEWATERING SYSTEMS AND ALTERNATIVES TO DEWATERING
There are probably as many potential dewatering systems as there are
underground coal mines, because the conditions that determine the best system
are to a large degree site specific. However, there are two basic types of
dewatering systems: vertical (conventional) wells drilled from the surface,
and angled drain holes drilled from the underground mine openings. These
dewatering systems are discussed in the following sections; in addition, some
alternatives to dewatering and to current mine drainage and water treatment
systems are suggested.
WELLS CONSTRUCTED FROM THE SURFACE
Vertical dewatering wells constructed from the surface either can be
drained by gravity into the mine openings or they can be pumped individually
to the surface. Parizek and others1* present schematic drawings of these
types of dewatering wells. Both systems can be effective in mine dewatering
provided that fracture zones and/or aquifers overlying and adjacent tp the
mine can be tapped.
The gravity-drained wells are probably most effective where a relatively
isotropic and more-or-less horizontal aquifer overlies the mine. Examples
Vould be a sandstone bed or channel, or a limestone bed, with considerable
areal extent, which both stores groundwater and transmits it to the mine.
Under these conditions, groundwater response to well drawdowns is reasonably
predictable and a regular well spacing can be used. The primary advantage
of this system is that individual well pumps are not required, therefore
capital and maintenance costs are restricted to those required for the pumps
which raise the collected water to the surface. In the case of drift mines
(mined up-dip), the system could be completely gravity drained, and there-
fore still more economical. Another advantage is the relatively small effect
that drainage wells have on the ground surface. Disturbance at the surface
is necessary only when wells are being constructed, and the only permanent
surface facilities would be at the mine water discharge point(s). The dis-
charge points could be located at the mine portal if all the drainage water
is collected underground; if, instead, water is pumped to the surface, a
single discharge 'point (a pumped well) could serve a group of drainage wells.
The primary disadvantage of wells drained by gravity into the mine is
lack of flexibility in location. Since they have to be drilled into mine
openings, they cannot be used to dewater in advance of mining. Also, where
fracture flow dominates mine inflows, it may be impossible to intercept in-
flows from outside of or beyond the edge of mine openings, since fractures
77
-------
above the mine may support only unsaturated flow, which cannot be intercepted
by wells. Further, in longwall mining, the panels are so large that it may
not be possible to penetrate important fracture zones to any significant de-
gree—especially fracture zone intersections, which may be particularly im-
portant to dewater. Finally, penetrating mine openings presents some safety
hazard because the breakthough could cause some amount of roof fall and there
will be uncontrolled water inflows (along with methane gas, if present) until
the drainage well is cased and hooked up to a header pipe. This may affect
the mine ventilation system and will cause some disruption of normal mining
activities.
In contrast,, pumped wells can be drilled through pillars to some depth
below the mine floors, or terminated at a safe distance above the mine roof;
they can be drilled in advance of mining and/or located around the periphery
of the mine. Aquifers can be tapped both above and below the mine, and there
is more flexibility in locating pumped wells where they can tap fracture
zones. The primary disadvantage, besides cost, would be that the facilities
needed on the surface are relatively hard to protect against weather and
vandalism. These facilities include powerlines to pumps and well discharge
pipes. However, to balance this disadvantage, the well discharges can be
collected and used on the surface for rural, municipal, or farm usage to
replace or supplement water supply from existing wells and springs whose flow
may have been affected by mining operations.
WELLS CONSTRUCTED UNDERGROUND
Standard mine air percussion drills can be used underground to drill ef-
fective drainage holes. These holes, about 6.4 cm (2.5 in.) in diameter,
could be drilled vertically into the roof or into the floor where mine in-
flows are greatest, or they could be angled into the roof to intersect the
the maximum density of fractures. Roof holes will drain by gravity, where-
as floor holes would be pumped and would act in similar fashion to a con-
struction well point system. The floor drains would be connected to a header
pipe, with one pump serving a series of wells to control water inflow through
the floor of the mine.
The roof holes would drain by gravity into a header pipe, which would
transfer water to a central pumping station where it is pumped to the surface
for discharge. These drainage holes can be designed either on a regular
spacing and pattern or concentrated where inflows are encountered when
driving entries and headings. Where fracture-dominated flow occurs, the
drain holes should be oriented to intersect a maximum number of fractures.
These holes can be drilled to arrays from both entries alongside longwall
panels, and angled toward the panel and a low-angle drain just above the
break line. Along the periphery of the mine, similar arrays of holes can be
used to intercept inflows from the sides.
For maximum efficiency, a specially equipped crew of miners probably
should be organized to drill drain holes. The drill should be mounted on
mobile equipment using conventional drills with 1.2 m (4.ft) rods, carbide
78
-------
bits, and standard 1.19 m /s (400 cfm) compressed air at 0.69 x 10^ Pa
(100 psi). To test the effectiveness of this system, trials can be made using
equipment normally available in the mine for drilling rock-bolt holes.
A disadvantage of wells drilled from the mine openings is that as water
inflows are encountered, they have to be handled individually in the conven-
tional manner, at least temporarily. Treatment, however, may not be required
if collection sumps and pumping stations for discharge to the surface are
constructed prior to mine advance. These pumping stations would have to be
over-designed to accommodate unanticipated inflows. Another disadvantage of
this system is that while dewatering can be accomplished in advance of mining
the longwall panels, it cannot be done in advance of driving headings and
entries. These drains may produce more methane than normal, which may affect
the ventilation systems however, this problem would be solved by proper col-
lection of the gas at the drain head, and the methane removal system might
serve wider applications than simply clearing the drains themselves.
There are two main technical advantages to this method of dewatering.
The first is its flexibility: a system can be designed at the start of
mining, based on exploration data, and easily adjusted as conditions dictate-
or drainage holes can be constructed on a trial-and-error basis as mining
advances. The second advantage is that the depth of the mine has little in-
fluence on the design of the system. The only adverse effect of depth on
costs would be the increased size of pumps, and thus the amount of energy
required to pump to the surface. However, the depth of the seam similarly
affects the costs of any of the dewatering systems. Thus, the merits of this
approach outweigh the disadvantages.
DESIGN APPROACH
The design of a dewatering system can only be as good as the concept on
which it is based. It is only as effective as the exploration, testing, and
interpretation of the data upon which the design is based. For a new mine
or extension of an existing mine, exploration of groundwater conditions
should be combined with coal exploration. In particular, the exploration
holes for coal can be designed for use in hydraulic testing and water quality
sampling and preserved for use in monitoring or actually incorporated into
the well dewatering system.
The surface geology of the area should be mapped with special emphasis on
linear features and joint patterns. Both high-level and low-level aerial
photography should be used. Existing wells and springs should be surveyed for
yield, water quality, and potentiometric levels. Surface geophysical surveys
should be performed to further delineate fracture zones and flow patterns. As
mentioned previously in Section 6, surface resistivity combined with surface
Potential streaming surveys may be the best geophysical techniques to consider.
Exploration drilling should be programmed to prove the existence, extent
and system of fracture zones interpreted from surface surveys, as well as to
explore the coal seam. Air-rotary drilling permits a rough estimate to be
made of the water-bearing properties of rocks encountered if the holes are
logged carefully and if the flow air-lifted from the holes is monitored as
79
-------
drilling progresses. Exploration holes should be cased sufficiently so they
can be used for water-level monitoring, and selected holes can be pump tested
for hydraulic properties and water quality.
With these data, potentiometric surfaces can be mapped, which, when com-
pared with fracture zones and pump tests, should indicate the areas of high
potential inflows to the mine. The interpretation can be enhanced by mathe-
matical modeling if the hydraulic data are reliable enough. A reasonable
basis should then be available for designing a dewatering system and its
technical and economic feasibility can be determined by comparing costs with
benefits.
ALTERNATIVES TO MINE DEWATERING
After the completion of this study, it became obvious that development of
AMD can be controlled or reduced by modifying existing mine drainage systems.
In an existing mine, before any new type of dewatering program or mine drain-
age treatment program is established, a total water quality survey of mine in-
flows should be performed. Further, a monitoring net should be established
for key parameters such as flow rate, pH, acidity, alkalinity, total iron,
and specific conductance, and monitoring should be performed on a continuing
basis until a basis for design is established. The monitoring program should
then be converted as necessary as part of the control program. These surveys
will establish the quantity and quality of the various underground flows and
will show which flows do or do not meet direct discharge standards. This
monitoring will indicate where water becomes degraded, if the quality is good
to begin with, or where inflows occur that require treatment. Unfortunately,
it is the experience of this study that monitoring of this type in most mines
is non-existent. However, it is exactly the type of information that is
needed before engineering decisions can be made in regard to the most effec-
tive solutions for water quality control.
With the above monitoring data, mine waters of different quality can be
separated and handled differently. Good quality water can be intercepted and
brought directly to the surface through bore holes and discharged without
treatment or piped through the mine to the shaft or slope. The water quality
information collected for the Lancashire No. 20 mine suggests that at least
60 percent of the total mine drainage could be intercepted within the mine and
discharged without treatment. In contrast, it is now transferred from sump
to sump where it degrades and is mixed with poor quality water from old work-
ings, so that all of the mine drainage requires treatment. For example, it
would have been a simple matter to: pump from the L-2 sump and G-ll sump into
the L-4 pond; drill a hole from the surface into the L-4 pond and install a
pump; then pump all of the Main G study area inflow to the surface and dis-
charge it into Laurel Lick Run.
Often, the main source of poor quality water is from old or abandoned
workings which may be difficult or too dangerous to enter. However, water
froty these workings can often be isolated and handled separately so that over-
all treatment costs are reduced even if the contributions from the active
portion of the mine are low in iron or acidity (or alkaline).
80
-------
Therefore, the experience developed during this study indicates that ad-
ditional research is needed to survey the opportunities for control of AMD by
determining water quality distribution throughout active coal mines. This
may result in more cost-effective measures, especially for existing mines,
than dewatering systems.
81
-------
SECTION 9
COST-EFFECTIVENESS OF DEWATERING
INTRODUCTION
Objectives
The previous sections have demonstrated that groundwater can be inter-
cepted by pumped surface wells before it enters an active underground coal
mine. It has also been shown that dewatering can result in smaller amounts
of AMD requiring treatment, since the water discharged will normally be of
the same quality as existing groundwater, unless a nearby contaminated source
such as flooded abandoned workings is also intercepted. The objectives of
this section are threefold: First, to establish cost comparisons between con-
ventional methods of handling and treating mine water and mine dewatering by
pumped surface wells. Second, to extend the well dewatering cost data to show
the influence on costs of variations in such elements as mine depth and water
quality, which differ from one mine to another and among different areas of
the same mine. Thirdly, to establish approximate costs for methods of de-
watering other than by pumped surface wells.
Study Conditions and Methods
In order to make the cost comparisons, two sets of basic cost data were
required. One set, pertaining to dewatering by pumped wells, was derivable
from the study project. To arrive at the second set of cost data, pertaining
to current methods of handling and treating mine water, and to cast the data
into a form that would allow valid comparisons with the well dewatering costs,
some research and mathematical manipulation were required.
The coal mining industry is highly competitive, and coal mine owners are
reluctant to allow operating and cost data to become public. Therefore there
is no official (or unofficial) industry standard of costs for in-mine water
handling or water treatment which might serve as a base of comparison with
costs established for the well dewatering project.
In fact, it may well be highly impractical to try to establish standard
figures for the industry since there is great variation in physical factors
from one coal mine to the next—size of mine, depth, water quality, engineer-
ing methods, etc.—so that costs can fluctuate greatly as a result of necessary
differences in operations. The actual range of operating costs for one ele-
ment of the mine water disposal system, say of water treatment alone, can be
great. Figure 33, derived from data collected by Skelly & Loy (1973)^ shows
a range of costs per cubic meter for water treatment ranging from about lc to
82
-------
sor
DC
U>
OBETHLEHEM MINES
NO. 58-A
908 m3/d
STUDY MINE
-11734 n3/d
10
D
0YOUNG 4 SON
681 m3/d
LOOMIS NO. 4
[21 802 B3/d
©WARWICK NO. 3
2271 m3/d
©1136 m3/d
O3407 i»3/d } *EST
QJO 220 m3/(| '
1136 m3/d }
3407 inVd >WEST VIRGINIA
^ 10 220 mVd-'
KOREA STRIP ^.WARWICK NO. 2
Ol5 140 m3/d WT1 446 m3/d
— 1 1 L
0123
O1136 m3/d \
Q3407 m3/d > WEST VIRGINIA UNIV.
OlO 220 «3/d )
VIRGINIA UNIV.
UNIV.
SLIPPERY ROCK CREEK
446 mj/d
1 | I
4 5 6
- OPERATION YELLOWBOY
OBET«I-EHEM
NO. 58-B
1136 m3/d
7 8 9
ACIDITY, mg/l (1000s)
DATA DERIVED FROM SKELLY AND LOY, 1973
Figure 33. Operating costs for hydrated lime treatment facilities.
-------
more than 20c depending on water quality. However, we do not know why the
acidity of the mine water is high or low, how much it is affected by opera-
tions, and whether the higher acidities are necessary and unavoidable. That
is to say, there are more physical variables influencing the final cost than
simply the recorded acidity level of the water. Furthermore, differences in
accounting practices can also contribute significantly to the range in costs
for treating approximately the same quality and quantity of water. Accounting
practices for all phases of mining industry activities are far from uniform,
as is illustrated in a study presented at the Second Symposium on Coal Manage-
ment Techniques and entitled "Survey of Accounting Principles, Procedures, and
Practices in the Coal Industry."^ xhe responses to the survey (Table 9)
illustrate clearly the lack of uniformity in expensing and capitalizing stan-
dard costs, and the study as a whole demonstrates further variations in
methods of depreciation, types of accounts, and the like.
TABLE 9. ACCOUNTING METHODS IN THE COAL INDUSTRY3
During Production
Item Capitalized Expensed Mixed Inventoried
Major repairs
Standby equipment
Replacement parts
Slurry ponds (0-3
years of useful life)
Construction changes
2
17
9
7
8
9 9
2
4
10
1
13
3. ;
Source: Monterey Coal Co., 1976.
Because of these variations in physical conditions, lack of uniformity in
accounting methods, and simple lack of published information, it is not pos-
sible to make valid comparisons of comprehensive and detailed costs between a
single operation and the industry as a whole, or even between two individual
operations. In evaluating the cost-effectiveness of well dewatering or of any
other methods of water disposal, the primary comparison has to be among the
methods available for the site in question. The considerations are both site-
specific in the geotechnical aspect and system-specific in the accounting
aspect. This is not to say that comparisons with data for other mines should
be totally avoided, for useful comparisons can be made as long as their limi-
tations are recognized.
The method of the present study, therefore, is to establish the costs of
current water disposal methods at 'the study mine and then to draw limited
comparisons between these data and costs for similar operations at other mines
insofar as these comparisons are useful. The extent of 'usefulness' of this
84
-------
data is its applicability to determining the cost-effectiveness of well de-
watering under conditions other than those at the study mine.
COSTS OF MINE WATER DRAINAGE AND TREATMENT - STUDY MINE
Accounting Methods at Study Mine
To make the required cost comparisons feasible, the accounting data for
mine water drainage and treatment at the Lancashire No, 20 mine must meet
three conditions: (1) costs for the mine under study must be clearly identi-
fied and not grouped with other mines of the company; (2) capitalization and
operating expenses must be clearly separated; and (3) the costs of water col-
lection and transfer must be separate from costs for water treatment to im-
prove quality. The Barnes and Tucker Company's accounting system meets these
requirements, with a few exceptions which do not significantly affect the re-
sults of this study. Costs of mine water drainage and treatment are separated
in the mine's accounting system by the method of cost accumulation, account-
ing, and payment. The two main categories of costs are capital investment
and annual operating expenses. It is apparent that under capital investment
all the costs for water treatment have been included (though these costs are
not clearly segregated from one another), except for land acquisition, defer-
red investment, contingency, and fee. From the information available, it can
be assumed that land acquisition costs were included under mine development
and can be so attributed. Deferred investment refers primarily to the dif-
ference between original cost and the cost of replacing equipment and con-
struction during the life cycle of the mine. These costs are therefore
included in capital investment. Contingency and fee are not itemized and are
introduced into this analysis as a factor of uncertainty. They generally run
to 10 percent of the initial capital investment cost. Because the mine has
been in operation for a number of years, the only capital costs for mine water
drainage are accounted for as deferred investment in equipment and construc-
tion projects. The equipment is limited to that required for the removal of
water from the mine.
The annual operating expense sections of the mine accounting system are
not ideal in a few respects. The division into direct, indirect, and fixed
expenses is not clear cut because the Company Service Department'includes
overhead in its charges to the mine operation for various services. The over-
head apparently includes payroll overhead, union welfare, and general and ad-
ministrative costs (G & A), the sum of which is about 90 percent. Taxes and
insurance have not been noted; for this analysis they are set at 2 percent of
the initial capital investment cost. Power and water costs are not tabulated
separately for the mine, preparation plant, and water treatment plant, and
this overall cost is in turn combined with similar costs for other mines as a
single, tabulated value. For purposes of this analysis, these values were
estimated. The methods of estimation are presented as part of the discussion
of these costs below.
Labor cost is not assigned to each mine for the two personnel employed to
oversee the operation and effectiveness of the water treatment of the four
mines owned by the company in the same local geographic area. However, this
85
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cost was assumed (probably with reasonable accuracy) to be the total yearly
salary of the two, with the addition of overhead at 90 percent, divided by
four.
An attempt was made to determine the completeness of cost entries, and to
identify joint costs and hidden costs. In most cases, missing cost entries
are recorded elsewhere in the accounting system as a joint cost. Joint costs
and missing entries were calculated using standard techniques. There are
certain hidden costs which result from entries being present but not complete,
particularly in connection with some types of labor costs. For example, only
pumper personnel charge time to the drainage account. In a wet mine, such as
this one, heavy influxes of water occasionally occur which require the work
of all the face section personnel for two or three days. The actual time
costs for one single such event can be equivalent to 10 to 30 percent of the
annual pumpers' labor cost. At least one such event occurred during the 1976
fiscal year (the period of the dewatering study). Another unrecorded labor
cost is the use of injured and other unassigned personnel to help in repair
of pumps on the surface. To compensate for these unrecorded costs, an error
of plus 20 percent of pumper labor cost is added to taxes and insurance (2
percent of initial capital investment) as an upper range of operating expenses
for drainage and water treatment.
Capital Investment and Annual Operating Costs
The water-related mine costs studied in this analysis are for the
company's 1976 fiscal year from October 1975 through September 1976. The de-
tailed costs, as provided by the company, are presented on Tables 10 through
13. Tables 10 and 11 itemize annual operating expenses for water collection
(drainage) and water treatment, respectively. Table 12 shows special work
orders for both drainage and water treatment over the 12 months of the
company's fiscal year. Table 13 presents the items capitalized as depreciable
assets for both drainage and water treatment.
Table 14 summarizes capital investment costs in which contingency and fee
are added and total depreciation to date is noted. The total investment cost
for drainage and water treatment is about $701,000, or $771,000 with contin-
gency and fee. The investment cost for drainage is about $310,000, or
$342,000 with contingency and fee. The investment cost for water treatment is
about $391,000, or $430,000 with contingency and fee.
Table 15 is a short summary of annual operating expenses in which esti-
mates for missing expense entries are added to the company's entries. The
items of depreciation for drainage and water treatment are minor entries.
Taxes, insurance, and unrecorded labor are calculated as operating costs which
are not recorded by the mining company. The percentage increase represented
by these costs is given. The unrecorded labor under drainage is estimated
from activities throughout the year above and below ground, including one in-
stance in which 2 sections of 12 men worked 3 days to control a particularly
heavy flow. This single incident accounts for 10 percent of the entry. The
accuracy of such entries can be questioned, yet an expense should be entered
86
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TABLE 10. SUMMARY OF DRAINAGE ANNUAL OPERATING EXPENSES
Labor
Fringe Benefits
Material
Direct Charges
CO
••J
Work Orders
Blanket Order
Blanket Order
Blanket Order
Blanket Order
Special Orders
Account Cost
No. Source
315 P/S. Journal
_
316 Acct. Pay (AP) Journal
AP & Supplyhouse
Std. Journal
#104
#112
#113
#114
Various
Description
Labor Cost for UMWA employees classified as pumpers
Estimated @ 90Z of labor
Outside charges for Material & Labor
Repairs to pumps
Purchase of New Pumps
Valves
Couplings, motors & other parts
Fiberglass pipe
Parts £r Labor & Overhead chgd. from Company Service Bept.
Routine Maint - Small Drainage Pump Motor
Routine Repair - #10 Labour pumps
Routine Repair - #23 Labour pumps
Routing Repair - All other small pumps
See Table 11
Total Material and Expense
Total Actual Cost Recorded - Underground Drainage
Actual
Costs Recorded in
Water Related Accts.
Detail Total
$92,523
$36,740
9,930
14,895
22,839
14.895 99,299
$15,088
-
1,478
159
6.807 23,532
122,831
$215,354
Allocation of
Costs Charged
Elsewhere
$83.270
-------
TABLE 11. SUMMARY OF WATER TREATMENT ANNUAL OPERATING EXPENSES
Account Cost
No. Source Description
Labor None No specific employees directly assigned - plant operated
by Service Department and Labor charged via Blanket
Work Order 0103, below
Fringe Benefits — — Charged via Work Order #103, see Table 11
00 Material
Direct Charges 394-4 Acct. Pay (AP) Outside charges for Material & Labor
Journal Lime
Pumps
Parts & Repairs to Pumps, etc.
Valves
Equipment Rental - Sludge Removal
Seeding & Mulching
Work Orders — Std. Journal Parts, Labor & Overhead chgd. from Company Service Dept.
Blanket Order #103 Operation 4 Routine Maint. of facilities
Special Orders Various
Total Actual Cost Recorded - Water Treatment
Actual
Costs Recorded in Allocation of
Water Related Accts. Costs Charged
Detail Total Elsewhere
$39,170
7,500
8,325
1,665
20,000
6,684 $83,344
$72,982
20, 131 93,163
$176,507
-------
TABLE 12. WORK ORDERS - WATER TREATMENT AND DRAINAGE - FISCAL 1976
oo
VO
Oct. Nov.
WATER TREATMENT A/C C 349-4
3W/0 #103 - Blanket Order $3,356 $3,840
#9747 - Construct pumping station to 502
convey cleaning plant water
to #2 Basin (iacl. ditch,
pipe, pump, foundation, etc.)
TOTAL $3.356 $4.342
DRAINAGE - A/C C 316
W/0 #9869 - Eepair 25 hp pump starter
49864 - Repair 40 hp electric motor
#9975 - Repair 300 hp Emco pump starters
- $9976 - Mount motor & pump on base
110114 - Repair 20 hp pump motor
#10134 - Repair 300 hp pump motor
#10135 - Repair 15 hp pump motor
#10184 - Repair Acme compressor
#10286 - Repair 50 hp trash pump motor
Sub Total
W/0 1104 - Blanket Order 183 1,451
#113 - Blanket Order 716 762
#114 - Blanket Order
TOTAL $ 899 $2,213
Dec. Jan. Feb. Mar. Apr. May June July Aug. Sept. Total
$3,902 $3,673 $3,352 $9,706 $4,962 $6,055 $9,251 $9,533 $6,877 $8,475 $72,982
8,665 10,828 186 20,181
$12.567 $14.501 $3.538 $9,706 $4,962 $6.055 $9,251 $9.533 $6.877 $8,475 $93.163
531 531
401 401
515 515
744 744
733 733
257 257
643 643
2,608 2,608
375 375
932 1,259 733 900 2,608 375 6,807
1,744 440 746 1,913 4,036 1,906 899 383 900 487 15,088
1,478
159 159
$2,676 $ 440 $2,005 $1,913 $4,928 $2,806 $3,507 $ 383 $1,275 $ 487 $23,532
All work orders include labor, fringe, and all departmental overhead
charged via departmental rate applied to actual hours.
-------
TABLE 13. SUMMARY OF ITEMS,CAPITALIZED AS DEPRECIABLE ASSETS - DRAINAGE AND WATER TREATMENT
Depreciation Purchase Accumulated
Life Date Cost Depreciation
UNDERGROUND DRAINAGE
Pumps - Various, incl. Deep-Well Pumps 10 yr. 1964 to 1976 $88,075 $24,054
Pipe - Fiberglass & PVC 10 yr. 1964 to 1976 69,885 25,625
Motors, Starters, etc. 10 yr. 1964 to 1976 39,271 9,164
Various Underground Projects
Such as: Sump areas, centrifugal pumps,
bore holes, bulkhead of old mine sec-
tion, etc. - incl. pipe pumps, etc. 20 yr. 1966 to 1968 113.249 40,867
Total Underground Drainage $310,480 $99,710
WATER TREATMENT FACILITIES
Old Treatment Facility 20 yr. 1967 26,552 21,319
New Treatment Facility 20 yr. 1971 to 1972 293,636 16,864
Hydraulic Rake System - Equipment 20 yr. 1975 to 1976 35,512 3,551
- Installation 20 yr. 1975 to 1976 35.052 1,753
Total Water Treatment Facility $390.752 $43.487
Net
Book
Value , Comments
Cost of Equipment, Only
Cost of Equipment, Only
Cost of Equipment, Only
Includes Vendor's cost for equipment,
contractors and company labor
and Service Dept. charges3
$210.770
Outside contractor and company labor
Includes cost of outside services for site
preparation, engineering, construction and
equipment plus charges for company
Service Dept.
Cost of Equipment, Only
Charges from company Service Dept.
$347,265
Charges from Company Service Dept. includes Labor, Fringe, and All Departmental Overhead.
-------
TABLE 14. SHORT SUMMARY OF CAPITAL INVESTMENT
Function
Drainage
Investment
Costa
$310,480
Accumulated
Depreciation
$99,710
Depreciation
Cycle
10 years
-pumps
20 years
-projects
Contingency/
Fee
$31,048
Drainage with
Contingency/Fee
341,528
Water
Treatment
390,752
43,487
20 years
39,075
Water Treat-
ment with
Contingency/Fee
429,827
Total Cost
$701,232
aAlthough data are presented to the nearest dollar, rounding to hundreds
or thousands of dollars would be appropriate for subtotals and totals, as
these numbers imply a precision not justified by the use of a single year's
costs, or by the method of accounting, or by the use of estimates where
data are missing.
Total Cost including Contingency/Fee: $771,355.
91
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TABLE 15. SHORT SUMMARY OF ANNUAL OPERATING EXPENSES
Type of Expense
Labor
Fringe Benefits
Material
Work Orders
Power and Water
Depreciation
Subtotal
Taxes/Insurance
@ 2% Capital Invest.
Unrecorded Labor
@ 20% Labor
Subtotal
Total
Drainage Cost
$92,523
83,270
99,299
23,532
196,430b
10,000°
$505,054
6,830
18,505
25,335e
$530,389
Water Treatment Cost
a
a
$83,344
93,163
__b
io,oooc
$186,507
8,597
14,250d
22,847f
$209,354
a
Assigned to 4 mines - costs not separated out.
The figure entered under drainage costs may be taken as for power only.
Costs of power for water treatment and of water for both drainage and
water treatment are within the error of the calculation of drainage
power cost. These estimates are not included in order to make the power
estimate even more conservative.
Q
Depreciation is a straight line apportionment of accumulated depreciation
back to initial date of installation of major investment for water treat-
ment (1972) and back to initial introduction and recording of pumps (1964)
d($30,000 + $27,000) f 4 mines.
f*
5.0% increase.
12.2% increase.
92
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to represent the fact that some charges accrued but are not recorded for the
activity. From Table 15, the total annual operating cost of water collection
and treatment is approximately $740,000.
The costs of electricity in operating the pumps are not separated from
other electrical costs in the operation of the mine. These energy costs were
estimated by obtaining the type and number of pumps with their full-load
horsepower rating. Allowance was then made for daily usage. The horsepower
was then converted to kilowatt hours and costs were established per kilowatt
hour. Finally, an estimate was made of percentage of full load at which the
pump is set. The resulting computed annual energy costs for the various
pumps are shown on Table 16. Also as shown on Table 16, the total annual mine
pumping cost is $157,158. The annual demand charge is not shown on Table 16,
but was calculated to be $39,272 and the total annual cost therefore is
$157,158 (from Table 16) plus $39,272, or $196,430. This total, shown on
Table 15, is 27 percent of the total cost for electricity of the study mine.
These figures confirm a gross estimate of 25 percent furnished by a knowledge-
able employee of the mining company.
The estimated average annual discharge from the mine based on pump rec-
ords is 13 400 m3/d (3.54 mgd). This value is only a gross measure because it
is based on a recording of the hours the pump was on and its rated operating
displacement. The pump flow from one of the pumps (F-14) was measured with a
flow meter over a period of 119 days, and the resultant average rate was
4400 m3/d (1.15 mgd). Similarly, the average value estimated from records for
this period was 5100 m3/d (1.34 mgd). As a result, the estimate based on the
F-14 pump records was about 14 percent too high. Therefore, the flow values
based on pump records were reduced to 11 700 m3/d (3.1 mgd), and this value
was used in cost calculations.
An important calculation for comparison purposes is the operating cost of
water collection (drainage) and water treatment. This cost is calculated
as follows:
A = 100 C (annual)
D x P
Where C is the annual operating cost
D is the number of operational days in the year
P is the daily pumping rate in m3
A is the water collection costs in C/m3
D „
A is the water treatment costs in C/mr
AD . $530 000 x 100 = 12>4 3
365 x 11,730
gal
365 x n>73° uy^iioS 8ai
^,$20^000^402 . *.**,„>, or
93
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TABLE 16. ANNUAL MINE COST OF ELECTRICITY FOR UNDERGROUND DRAINAGE
Flow
Point
F-14
F-l
K-7
Slant
Slant
Main-C
G-5
F-17
G-7
G-9
K-12
Fresh H20
G-10
G-ll
F-13
F-17
F-20
K-7 Belt
Main H
F-20
Faces
Type of Pump
1500 gpm Gould
1500 gpm Gould
2500 gpm Gould
1000 gpm Wilson Snyder
2500 gpn Hazelton
1000 gpm Flood City
1000 gpm Flood City
1000 gpm Flood City
1000 gpm Hydramatic
1000 gpm Hydramatic
1000 gpm Hydramatic
1000 gpm Gould
150 gpm 23 Labour
150 gpm 23 Labour
150 gpm 23 Labour
150 gpm 23 Labour
150 gpm 23 Labour
150 gpm 23 Labour
50 gpm 10 Labour
50 gpa 10 Labour
50 gpm 10 Labour
Totals
No. of
Pumps
1
1C
1
1
1
1
1
1
1
1C
1
1
1
1
1
1
2
1
1
1
40
64 C
Power per
Unit (hp)
300
300
300
125
100
50
50
50
SO
50
50
50
15
15
15
15
15
15
5
5
5
2140°
Total3
Load (hp)
255.0
255.0
255.0
106.3
85.0
42.5
42.5
42.5
42.5
42.5
42.5
42.5
12.8
12.8
12.8
12.8
25.5
12.8
4.3
4.3
170.0
Period of
Full Load (h/d)
,24
24
5
24
24
2
3
8
12
12
12
12
24
12
24
16
24
24
8
8
12
Energy Consumption
kWh/d
4,565.5
4,565.5
951.2
1,903.2
1,521.8
63.4
99.1
253.6
380.5
380.5
380.5
380.5
229.2
114.6
229.2
152.8
456.6
229.2
25.7
25.7
1,521.8
18,430.1
Daily Costb
$100.44
100.44
20.93
41.87
33.48
1.39
2.18
5.58
8.37
8.37
8.37
8.37
5.04
2.52
5.04
3.36
10.05
5.04
0.57
0.57
33.48
$405.46
Annual Cost
$36,660.60
45,825.00
7,639.45
15,282.55
12,220.20
507 . 35
795.70
2,036.70
3,055.05
3,055.05
3,055.05
3,055.05
1,839.60
919.80
1,839,60
1,226.40
3,668.25
1,839.60
208,05
208 . 05
12,220.20
$157,158.10
% Full load = 0.85 is included in this calculation.
Total rate per kWh is $0.007 basic plus $0.015 fuel adjusted, for a total of $0,022.
There is a second pump on standby. Figures for occasional use of this pump are included in total.
-------
A summary of totals of annual operating expenses and capital investment
for fiscal 1976 is given in Table 17. The water treatment required greater
capital investment and underground drainage required greater annual expense.
TABLE 17. MINE COST SUMMARY
Type of Cost
Water
Collection
Water
Treatment
Total
Annual Operating Expenses
Capital Investment
$530,000
$342,000
$209,000
$430,000
$739,000
$772,000
Major Cost Contributors
In this section, the major contributors to the costs of the study mine's
current water disposal system are presented, as far as they can be discrim-
inated. Contributors to drainage costs are discussed first, and contributors
to treatment costs second. The separation between operating expenses and
capital investment is maintained in each case.
The cost contributors in annual operating expenses for underground drain-
age are presented in Table 18. There are two functional categories of under-
ground drainage: (1) collection and transfer of water to the bore hole, and
(2) pumping to the surface. A key separation is made between these two sets
of cost contributors to permit comparison to the dewatering major cost con-
tributors presented later. Although costs are not recorded according to this
categorization, approximate estimates can be made. The procedure is to cal-
culate the percentage of each cost contributor (e.g. power,) that is used
either to collect and transfer water or to pump water to the surface.
TABLE 18. UNDERGROUND DRAINAGE ANNUAL OPERATING EXPENSES -
COST CONTRIBUTORS
Deprec.
Item Power Labor Material Orders and Taxes Total
Water Collection
and Transfer
Pumping to Surface
Total
% of Total
$92,702 184,583 99,299 23,532 8,415 $408,531
$103,728 9,715 0 0 8,415 $121,858
$196,430 194,298 99,299 23,532 16,830 $530,389a
37 36 18.7 4.4 3.2 100%
See Table 15.
95
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Power costs are the main contributor (37 percent) to the annual operating
expenses shown on Table 18. In calculating power costs, Table 16 is key be-
cause it presents all the pumps with their purpose noted. Three pumps (F-l,
F-14, Main C) are used to discharge water from the mine to the surface.
Using the power costs of these three pumps on Table 16, the cost of electricity
for pumping to the surface is computed to be 52.8 percent of the total cost of
electricity on that table:
$36.660 + $45,825 + $507 _ $82,992 _ R0
$157,158 $157,158 ~ ^-0/°
Using this percentage to prorate the annual demand charge and adding the re-
sult to the energy .charge, the total annual power cost to pump the water to
the surface is:
$82,992 + (0.528 x $39,272) = $103,728 (as shown on Table 18)
The percentage of the energy charge for collection and transfer is 47.2 per-
cent (100 percent less 52.8 percent) and thus the total power cost for
collection and transfer is:
($157,158 - $82,992) + (0.472 x $39,272) = $74,166 + $18,536 = $92,702
(as shown on Table 18)
The next major contributor of underground drainage annual expenses is
labor (36 percent). In this case, however, the operation and maintenance of
the water collection and transfer system is far more extensive than simply the
maintenance of three mine discharge pumps. The number of pumps is greater
and the distance traveled for maintaining the operation is far greater. It is
estimated that the labor cost to pump water to the surface is about 5 percent
of the total labor cost for underground drainage or $9,715, leaving $184,583
to cover labor costs for collection and transfer of water.
The next largest cost contributor is material (18.7 percent). All of
the cost is allocated to water collection and transfer because the pumps and
piping are capitalized. Thus, material is set at 100 percent water collection
and transfer or $99,299. The smaller cost contributors are work orders (4.4
percent), and depreciation and taxes (3.2 percent). Work orders relate total-
ly to water collection and transfer costs, while depreciation and taxes are
allocated at 50 percent to each of the two functions.
Of the two sets of functional cost contributors in total underground
drainage annual expenses (Table 18), water collection and transfer is by far
the larger, representing 77 percent ($408,531) of the total cost, while pump-
ing to the surface costs only 23 percent of the total ($121,858).
The underground drainage capital investment costs (Table 17) could not be
separated into these two functional cost contributors. The cost contributors
to drainage capitalization are primarily in equipment and piping. Equipment
accounts for at least two-thirds of the cost. Drilling of boreholes, con-
struction of bulkheads, and construction of sump areas are the source of the
remaining one-third. Labor is included in the construction projects only.
96
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The annual operating expenses of water treatment are separable by the
usual divisions of labor, depreciation, repair, etc. The expense tabulations
provided by the mine company are presented in Table 19. The major cost con-
tributor in this table is the item "labor, material, and repair" because it
cannot be broken down any further. "Repair" includes maintenance and operation
as well as actual repair. "Lime" is separated from "material" and is the next
largest cost contributor. The third largest is "sludge disposal". This entry
includes only the truck rental, and not labor or any other source of expense.
"Miscellaneous" includes emission tests and taxes (Table 15).
TABLE 19. WATER TREATMENT ANNUAL OPERATING EXPENSES
COST CONTRIBUTORS (1976 FISCAL YEAR)
Labor, Material,
and Repair
Cost ($) 124,923
% Total 59.7
Lime
39,170
18.7
Sludge ,
Deprec. Disposal u
10,000
4.8
20,000
9.6
Misc .a
15,261
7.3
Total0
$209,354.
100. l%d
Derived from information shown on Tables 11 and 15.
bSee Table 11.
CSee Table 15.
Rounding error.
The capital costs for water treatment could not be separated into func-
tional cost contributors. The various parts of the treatment plant such as
thickeners, sludge ponds, etc., could not be identifiedrfrom the cost data
provided by the mining company.
Summary .
In the preceding discussion, water-handling costs at the study mine have
been analyzed to identify cost factors that will permit comparison with other
systems in the industry and with the pilot dewatering operation. The analysis
is summarized in chart form on Figure 34. The water-handling costs have been
divided first into the conventional categories of capital investment and annual
operating expenses. Each of these categories is then subdivided into costs
for underground drainage (collection and transport of water underground) and
for water treatment. This second division is key to the later comparisons,
since the water treatment costs become a means of comparing the study mine to
the industry, and also constitute the primary savings to be anticipated in a
well dewatering system. Subsequently, cost contributors to annual operating
expenses for both underground drainage and water; treatment.have been identified
or (where necessary and possible) estimated, again to facilitate later
97
-------
CAPITAL
INVESTMENT
$772,000
STUDY MINE
WATER
HANDLING
COSTS
UNDERGROUND
DRAINAGE
$530,000
WATER
TREATMENT
$209,000
COLLECTION
AND
TRANSFER
$408,000
PUMPING
TO
SURFACE
$122,000
POWER
$92,702
MATERIAL
$99,299
LABOR
$184,583
WORK ORDERS
$23,532
DEPR. & TAXES
$8.415
POWER
$103.728
LABOR
$9.715
OEPR. & TAXES
$8,415
LABOR, MATERIAL
AND REPAIR
$124.923
Til
$39,170
DEPRECIATION
$10.000
SLUDGE DISP.
$20,000
MISC.
$15,261
Figure 34. Summary of cost analysis.
98
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comparisons. The available data on capital Investment were not appropriate
for identification of cost contributors to capitalization of either under-
ground drainage or water treatment.
INDUSTRY COSTS
As noted earlier in this section, data on costs in the coal mining indus-
try are infrequent. There appear to be no data published on costs of under-
ground drainage, and therefore we have to assume that costs at the study mine
are somewhat representative for the industry.
On the other hand, the question of treatment of AMD has been a focus of
public attention as well as a concern of the industry for some 10 to 15 years,
and reliable cost data' have emerged, especially from the Operation Yellowboy
Program, reported by Charmbury et al., 1967 . In this program conducted
during the 1960's at the University of Virginia Mine No. 1, a lime treatment
facility was operated under study conditions, and consistent accounting
methods were used to determine capital and operating costs for various water
quality conditions and plant capacities. These cost data have been used by
Holland et al.^ with reference to treatment plant design, and Holland's work
has been used by Shumate et al. to develop predictive models for hydrated
lime neutralisation capital investment and annual operating costs. Cyrus Wm.
Rice & Company-'-^ also use these data in evaluating the economics of lime neu-
tralization as a method of water treatment. Doyle, Bhatt, and Rapp draw on
all of these earlier works (except Shumate) in their review of water treatment
methods, and they present several tables and charts which provide a useful
industry context for the study mine. All cost elements in Doyle et al. are
stated in 1972 dollars. Since study mine costs are based on 1976 accounts,
study mine costs have been converted to 1972 dollars in the following compar-
ative discussions using the construction contract cost index presented below.
Construction Contract Cost Index from 1967-1976
Year Index
1967 100
1968 113.2
1969 123.7
1970 123.1
1971 145.4
1972 165.3
1973 179.5
1974 169.7
1975 166.0
1976 168.1
Tables 20 and 21 together present a broad view of treatment costs in the
industry. They show capital and operating costs, respectively, for ten treat-
ment operations and also for the Operation Yellowboy study plant. Table 22,
based on Operation Yellowboy data, shows a range of cost contributors to costs
of hydrated lime treatment at different levels of acidity and plant capacity.
Figure 35 and 36, based on the same reference data, plot capital costs and
99
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TABLE 20. SUMMARY OF CAPITAL COSTS - CONVENTIONAL LIME
NEUTRALIZATION PROCESS3
Mine
Bethlehem Mines Co.
No. 58-A
Bethlehem Mines Co.
No. 58-B
Young & Son
Morea Strip
Blue Coal Corp.
Loorais No . 4
Duquesne Light Co.
Warwick No. 3
West Virginia Univer-
sity School of Mines
Mine No. 1
(Operation Yellowboy)
Duquesne Light Co.
Warwick No. 2
Commonwealth of Pa.
Slippery Rock Creek
Treatment Plant
Rausch Creek Mine
Drainage Plant
Mountaineer Coal Co .
Design
Flow Rate
m3/Day
908
1 136
681
15 140
21 802
2 271
1 136
1 136
1 136
3 407
3 407
3 407
10 220
10 220
10 220
11 446
11 446
37 850
2 725
Total
Acidity
4 080
8 150
770
190
560
1 250
3 500
1 400
650
3 500
1 400
650
3 500
1 400
650
1 560
240
250
Capital
Total
Cost
$ 347,200
423,300
229,900
657,400
1,094,000
229,700
582,000
750,000
1,747,380
120,000
Costs
$/m3
$382.38
372.53
337.59
43.42
50.18
101.14
—
— _
50.85
65.53
46.17
44.04
Source: Skelly and Loy, 1973.
100
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TABLE 21. SUMMARY OF OPERATING COSTS - CONVENTIONAL LIME
NEUTRALIZATION PROCESS3
Operating Costs
Mine
Bethlehem Mines Co.
No. 58-A
Bethlehem Mines Co.
No. 58-B
Young & Son
Morea Strip
Blue Coal Corp.
Loomis No. 4
Duquesne Light Co.
Warwick No. 3
West Virginia Univer-
sity School of Mines
Mine No. 1
(Operation Yellowboy)
Duquesne Light Co.
Warwick No. 2
Commonwealth of Pa.
Slippery Rock Creek
Treatment Plant
Rausch Creek Mine
Drainage Plant
Mountaineer Coal Co.
Design
Flow Rate
m3/Day
908
1 136
681
15 140
21 802
2 271
1 136
1 136
1 136
3 407
3 407
3 407
10 220
10 220
10 220
11 446
11 446
37 850
2 725
Total
Acidity
mg/£
4 080
8 150
770
190
560
1 250
3 500
1 400
650
3 500
1 400
650
3 500
1 400
650
1 560
240
250
Annual
Cost
$ 95,250
140,000
47,400
126,571
475,000
117,500
68,448
44,236
30,223
172,463
108,405
73,913
477,968
290,723
196,607
109,715
51,000
c/lOOOmli
per
mg/£ acidity
7.0
4.1
24.8
12.1
10.7
11.3
4.7
7.6
11.2
4.0
6.2
9.1
3.7
5.6
8.1
3.2
5.1
Source: Skelly and Loy, 1973.
101
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TABLE 22. COMPARISON OF COST CONTRIBUTORS: STUDY TREATMENT FACILITY
VS. OPERATION YELLOWBOY RESULTS
Labor, Material
Study Treatment Facility3 and Repair
100 mg/£ acidity - 3.7 mgd capacity
% of Annual Operating Expenses 59.7
Operation Yellowboy Facility Labor Repair
150 mg/£ acidity - 2.7 agd capacity 15-3 4.3
19.6
Sludge
Lime Depreciation Disposal Misc.
18.7 4.8 9.6 7.3
Sludge
Lime Plant Cost Disposal Misc.
19.7 43.6 4.3 12.3
Total
100. 1C
Total
100
From Table 19.
Developed from costs data extrapolated from Table 20.
Rounding error.
-------
CO
u
STUDY MINE
TREATMENT
PLANT
BASED ON DOYLE. BHATT, AND RAPP (1974)
0.2 0.3 0.5
1.0 2.0 3.0
CAPACITY (MGO)
10.0
Figure 35. Capital cost
hydrated lime
of treatment plant vs. plant capacity
treatment with sludge disposal.
103
-------
10.0
0.1
0.1
1.0
CAPACITY (MGD)
10.0
1-HIGHLY ACIDIC (WITH SLUDGE DISPOSAL)
2-HIGHLY ACIDIC (WITHOUT SLUDGE DISPOSAL)
3-MOOERATELY ACIDIC (WITH SLUDGE DISPOSAL)
4-MODERATELY ACIDIC (WITHOUT SLUDGE DISPOSAL)
5-WEAKLY ACIDIC (WITH SLUDGE DISPOSAL)
6-WEAKLY ACIDIC (WITHOUT SLUDGE DISPOSAL)
IRON (mg/l) ACIDITY (mg/l)
HIGHLY ACIDIC
MODERATELY ACIDIC
WEAKLY ACIDIC
900-1200
600-700
322
2600-4800
1400
600-700
BASED ON DOYLE, BHATT, AND RAPP (1974)
Figure 36. Total operating cost (including capital costs)
• hydrated lime treatment.
104
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operating costs, respectively, against plant capacity for different acidity
levels. On these figures, data for the study mine have been added to the
reference data.
On Figure 35, the solid-line plots for 200 mg/£ through 5000 mg/£ acidity
are original Operation Yellowboy data. At the study mine the exact quality of
the input water to the treatment plant is not known. It is a mixture of ap-
proximately 6700 m3/d (1.78 mgd) of 105 mg/£ acidity and 50 mg/£ total iron
with 5000 m3/d (1.32 mgd) of 15 mg/£ alkalinity and 12 mg/£ total iron, but
the proportions of the mix vary considerably and nonpredictably. It may be
conservatively assumed that the mixed input is 11 700 nr*/d (3.1 mgd) of 100
mg/£ acidity and 34 mg/£ total iron. A dashed line representing study facility
flow of 100 mg/£ acidity has been added to Figure 35 using a ratio method of
extrapolation. This level of acidity is considerably lower than the low end
of the range of reference data (200 mg/£).
The study mine's design capacity versus capital cost has also been plot-
ted (circle) on Figure 35. The construction cost for the study mine treatment
plant (maximum capacity of 19 000 m3/d—5 mgd) is $430,000 (Table 17), or
$423,000 in 1972 dollars. Since the acidity level of the water input to the
plant was very conservatively estimated at about 100 mg/£ (dashed line) and
the expectable acidity according to the reference data is about 89 mg/£, then
the design cost of the plant appears to be within the range of typical values,
and perhaps on the low side.
Figure 35 serves to compare the operating costs of the study mine treat-
ment plant with others in the industry, as represented by Operation Yellowboy
models. Treatment costs (including sludge disposal) at the study mine are
4.9 C/m3 (18.5
-------
total percentage for labor, materials, and repair. It may be noted that if
the percent costs for these items at the study mine facility are added to the
percentage for depreciation, the total (64.5 percent) is approximately com-
parable to the total of percentages for Operation Yellowboy labor, repair, and
plant cost (63.2 percent). Sludge disposal cost for the study facility is
about double that in the reference data because of truck rental expenses.
Miscellaneous costs, on the other hand, are 75 percent higher for the more
sophisticated Operation Yellowboy facility since these costs include water
quality monitoring and other testing. Though such variations in the relative
percentage components of annual operating costs are relatively large, in the
long run they apparently offset one another.
It may be concluded that water treatment costs at the study mine facility
(in terms of both capital outlay and yearly operating expenses) are approxi-
mately representative of industry costs for a plant of its capacity and for
water of this acidity level.
COSTS OF MINE DEWATERING USING PUMPED WELLS
The costs of mine dewatering using pumped wells are separated into oper-
ating expenses and capital investment. These are further segregated and
costed for electrical service (at the same rates as charged at the study mine),
maintenance labor and overhead, operation labor and overhead, and depreciation
of capital investment. All capital costs are depreciated over ten years
(straight line). Included as capital investment are maintenance parts, well
construction, pumping facilities, and engineering.
Three dewatering wells 152 m (500 ft) deep are costed, which represents
the number and the approximate depths of the wells in the pilot dewatering
operation. These costs are subsequently used in the comparison with the mine
cost reduction, represented by the amount of water intercepted before entering
the mine. The major cost contributors are well construction (capital invest-
ment) and electrical service (operating expenses). The cost items contributing
to capital costs are listed on Table 23 without identified subcosts. The op-
erating costs are presented on Table 24.
To illustrate the effect of well depth on costs of dewatering, the capital
and operating costs for one well were manipulated to correlate with well depths
between 30 m (100 ft) and 305 m (1,000 ft). These data are graphically pres-
ented in Figure 37 (less $2,500 for access rights). The increased slope for
capital investment after the 132 m (400 ft) depth is due to the change in well
and casing diameters. The well and casing diameters are increased from 20 cm
to 25 cm (8 in. to 10 in.) and 15 cm to 20 cm (6 in. to 8 in.) respectively.
Operating costs increase linearly with increased depth. To dewater dif-
ferent mines, different numbers of wells of various depths would be required.
Figure 38 graphically presents the variation in annual operating costs with
number of wells for well depths of 30.5 m (100 ft), 61 m (200 ft), 122 m
(400 ft), 152 m (500 ft), 183 m (600 ft), 244 m (800 ft), and 305 m (1,000 ft).
Figure 39 shows capital costs versus number of wells and well depths. The
106
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TABLE 23. CAPITAL COSTS FOR THREE WELLS 152 m (500 ft) DEEP
Cost Item Cost
Well Construction (total cost) $34,269
Fringe, overhead, administration
Drilling
Well Development
Casing (material and labor)
Expenses at site
Mobilization
Access Rights 2,500
Pumping Facilities (total cost) 10,794
Pump
Discharge pipe, cable
Pump installation (labor)
Wire pump to electrical source (labor)
Engineering (15% of above less access rights) 6,759
Labor
Fringe, overhead, administration
Total $54,322
costs increase directly in proportion to the number of wells. The major cost
contributors (well construction in capital costs, and electrical service in
operating costs) maintain their percentage of costs with an Increase in
number of wells.
COST-EFFECTIVENESS COMPARISONS
Costs have been developed in the foregoing discussions for the present
system of mine drainage and treatment, and for dewatering from pumped wells.
In this segment, cost comparisons are presented for the effectiveness of the
Pilot dewatering operation and for projected full scale dewatering of the
presently active part of the mine.
107
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TABLE 24. ANNUAL OPERATING COSTS FOR THREE WELLS 152 m (500 ft) DEEP
Cost Item Cost
Electrical Service $30,606
Maintenance 3,870
Labor
Fringe, overhead, administration
Operation 5,832
Labor
Fringe, overhead, administration
Depreciation of Capital Costs 5,184
(10-year life, straight line, less
access rights)
Total $45,492
Cost-Effectiveness Based on Pilot Dewatering (3 wells)
A comparison is made between the reductions in both annual operating ex-
penses and capital investment costs vs. the cost of pilot dewatering (3 wells)
In the existing water handling system, mine drainage and water treatment both
contribute to annual operating expenses and to capital investment; well de-
watering includes operational and investment costs of wells only, because no
treatment is required for the water discharged from the wells. The method
used is to estimate the reduction of mine drainage and treatment costs from
the ratio of the inflow reduction to the average annual mine drainage.
The average mine inflow reduction in the Main G study area was 2.33 t/s
(37 gpm), or 200 m /d (0.053 mgd) during the pilot dewatering period, as
opposed to an average annual flow of 11 700 nrVd (3.1 mgd) for mine drainage.
The reduction is 1.7 percent of the total mine drainage. This percentage is
projected to an annual reduction and applied to the values on Table 17 (last
column) for annual operating expenses and capital investment; the resulting
reductions are then compared with dewatering costs (from Tables 23 and 24).
108
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40
X
16
vt
o
o
-- CAPITAL INVESTMENT
— ANNUAL OPERATING EXPENSES
I
I
|
I
I
I
I
30.0
(98.4)
60.0
(196.B)
90.0
(295.2)
120.0
(393.6)
150.0
(492.0)
180.0
(590.4)
DEPTH OF WELL IN METERS (Feet)
210.0 240.0 270.0 300.0
(688.8) (787.2) (885.6) (984.0)
Figure 37. Operating and capital costs per well vs. well depth.
-------
1 000 r-
o
a
*>
CO
o
o
CD
400 -
200 -
304.8 m
(1 000 ft)
30
NUMBER OF WELLS
Figure 38. Annual operating costs vs. number of wells and well depth.
-------
000 -
NUMBER OF WELLS
Figure 39. Capital investment vs. number of wells and well depth.
304.8 m
(1000 it}
243.4 m
(800 ft)
182.9 m
(600 ft)
-------
These costs and the comparisons are shown on Table 25. The reductions of
drainage and treatment costs based on the 45 percent average well-effective-
ness ratio of reduction of mine inflow to total well yield obtained during
the pilot dewatering period, are $12,600 in annual operating costs and
$13,100 in capital costs. These cost reductions are then divided into oper-
ating and capital costs for the three dewatering wells, resulting in cost
ratios of 3.6 and 4.1 respectively. This indicates that the three-well de-
watering system, at 45 percent effectiveness, is 3.6 times greater in annual
operating costs and 4.1 times greater in capital costs than the corresponding
savings it produces.
As explained previously in Section 7, the well-effectiveness is estimated
to range between 50 and 80 percent depending on pumping time and recharge
events. Theoretically, it should approach 80 percent effectiveness after 120
days of pumping providing there are no major recharge events and no recharge
boundaries are intercepted. These levels of 50 and 80 percent effectiveness
would result in a decrease in mine drainage by 1.9 and 3.1 percent respectively
or 220 m3/d (0.059 mgd) and 360 m3/d (0.095 mgd). As shown on Table 25, at
80 percent well-effectiveness, the three dewatering wells are 2.0 times more
costly in terms of annual operating costs and 2.3 more costly than capital
costs of the existing system.
TABLE 25. PILOT DEWATERING COSTS
VS. MINE DRAINAGE AND TREATMENT COSTS
Item
Annual
Operating Capital
Costs Investment
Total Drainage and Treatment Costs ($1,000) 739
Dewatering Costs-3 pilot dewatering wells ($1,000) 45
Reduction of Drainage and Treatment Costs ($1,000)
45% well effectiveness (pilot dewatering period) 12.6
50% well effectiveness 14.0
80% well effectiveness 22.9
772
54
13.1
14.7
23.9
Ratios of Dewatering Costs to Cost Reductions (3 wells)
45% well effectiveness 4.7
50% well effectiveness 3.2
80% well effectiveness 2.0
4.1
3.7
2.3
112
-------
In the same manner, costs can be projected for dewatering the entire
active area of the mine, as shown on Table 26. The drainage from the active
area is estimated to be 60 percent of the total mine drainage or 7000 m^/d
(1.86 mgd). Assuming an average well yield of 1.9 £/s (30 gpm) at 80 percent
effectiveness, it would take 54 wells to completely dewater the active area.
From Figure 37, the annual operating cost would be $820,000 and from Figure 3BY
the capital cost would be $930,000. Dewatering with these 54 wells would re-
duce drainage and treatment costs by $443,000 in annual operating costs and
$463,000 in capital costs. The cost of dewatering with these wells would be
1.9 times more expensive and 2.0 times more expensive than the reductions in
operating costs and capital costs respectively. Similarly, Table 26 shows
that if the well-effectiveness is only 50 percent, 86 wells would be required
and they would cost 2.9 and 3.2 times more than the reductions in annual op-
erating costs and capital costs respectively. These ratios are slightly lower
than the ratios for pilot dewatering (Table 25) because they reflect a slight-
ly higher average well yield of 1.9 l/s (30 gpm) than experienced in the pilot
operation. This higher well yield should be easily attainable with better
well design, well siting, and pump selection. The costs of dewatering reflect
these improvements.
These calculations all indicate that dewatering with vertical pumping
wells would be at least twice as costly as current mine drainage and water
treatment at this mine and that dewatering apparently would not be a cost-
effective means of controlling AMD in this particular situation. However,
this does not mean that a dewatering system of this type is not economically
TABLE 26. ACTIVE MINE AREA DEWATERING COSTS
VS. MINE DRAINAGE AND TREATMENT COSTS
— - — -1-Annual
Operating Capital
Item Costs Investment
Total Drainage and Treatment Costs ($1,000) 739 772
Dewatering Costs ($1,000)
80% well effectiveness w/54 wells 820 930
50% well effectiveness w/86 wells 1,290 1,470
Reduction of Drainage and Treatment Costs 443 463
for 60% of total mine drainage handled
by dewatering ($1,000)
Ratios of Dewatering Costs to Cost Reductions
80% well effectiveness 1'9 2<0
50% well effectiveness 2'9 3t2
113
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feasible. If fracture zones can be effectively located by exploration, aver-
age yield of dewatering wells can be greatly increased. An increase of three
to four times the average yield (1.9 £/s—30 gpm) used in this study or 5.7
to 7.6 £/s (90 to 120 gpm) per well would make this type of system cost-
effective at the study mine, both in terms of operating costs and capital
costs. There are other benefits to be derived from mine dewatering that
should be considered. Loss of mine production from unstable roofs and sudden
large inflows of water can be reduced or possibly eliminated by dewatering.
In some mines, increased production could therefore be the largest single
benefit derived from mine dewatering.
Other potential benefits may include more opportunities to use the water
pumped from the dewatering system by overlying and adjacent communities or
land owners. Also, the water discharged from wells is better in overall
quality (at least in the case of this study) than treated water. It has less
total dissolved solids and sulfates than the effluent from the treatment
plant. Although it would be difficult to place a dollar value on this ben-
efit, it could have only positive effects on agricultural and consumer water
applications. There are probably many other benefits to the mining operation.
All of these potential benefits will vary widely from mine to mine, and the
economic feasibility of mine dewatering must be considered separately for each
case.
Effect of Water Quality on Dewatering Costs
From the foregoing discussions, it is apparent that dewatering would be
cost effective if the quality of mine drainage was so poor that the cost of
treatment would offset dewatering costs. By using the ratios developed for
full scale dewatering shown on Table 26 for capital costs, the capital treat-
ment costs and annual treatment operating expenses required to break even with
full scale dewatering can be calculated. The mine drainage costs are assumed
to remain the same and it is assumed the well discharges would not require
treatment. The break-even capital investment for treatment using a well-
effectiveness of 80 percent would equal (2.0 x $772,000) - $342,000 or
$1,202,000. Plotting this value on Figure 33 for a capacity of 11 700 m3/d
(3.1 mgd) gives an acidity of about 500 rngAt. The corresponding break-even
operating treatment expense would be (1.9 x $739,000) - $530,000 or $874,000.
This value equals 7.4 C/m3 (28 C/1,000 gal), and when plotted on Figure 34
falls slightly under the weakly acid curves (600-700 mg/£).
For a well-effectiveness of 50 percent, the break-even capital investment
cost is (3.2 x $772,000) - $530,000 or about $2,128,000. Plotted on Figure 35
for a capacity of 11 700 m3/d (3.1 mgd), this value results in an acidity of
about 1100 mg/£. The break-even operating cost at 50 percent well-effective-
ness is ^2.9 x $739,000) - $430,000, or about $1,613,000, which reduces to
13.7 C/m (52 c/1,000 gal). This value is closest to the moderately acidic
curves on Figure 36. These calculations indicate that dewatering would be
cost-effective if the acidity ranges between 500 and 700 mg/£ and the de-
watering wells are 80 percent effective in intercepting mine inflows. If the
114
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dewatering wells are 50 percent effective, the acidity would have to be in
the 1000 to 1500 mg/£ range. Again, if other benefits such as roof stabili-
zation are of importance, it may be economically feasible to dewater with
substantially lower acidities than the ranges estimated here.
Effect of Mine Depth on Dewatering Costs
The depth of the mine would affect the costs of dewatering primarily due
to well depth and pump size requirements (capital costs) and energy require-
ments (annual operating expenses). Mine costs would also change because of
different pump sizes required at the mine exit stations (capital cost) and
because of different energy requirements (operating expenses). Projections
of data developed herein for the Lancashire No. 20 mine revealed that dewater-
ing would not be cost-effective unless the mine was less than an average of
67 m (220 ft) deep when the average dewatering well-effectiveness was 80 per-
cent. For a dewatering well-effectiveness of 50 percent, the mine has to
average less than 30 m (100 ft) deep. The divergence between dewatering
costs and reductions of drainage and treatment costs is rapid with depth, and
other benefits must be found to make up the differences. It does suggest
that the shallower the mine, the more serious the consideration that should
be given to mine dewatering for water quality control.
Effect of Dewatering System Design on Costs
There are other dewatering systems that can be designed which appear to
be less costly than the system used during this study. Vertical wells can be
drilled from the surface into mine openings and the water in each well can be
allowed to drain by gravity into a header pipe. The water is then transferred
to a central pumping plant from which it is pumped to the surface for dis-
charge. The capital costs of this system would be about 15 to 20 percent less
because only one pump is used for a group of wells (or possibly all wells) and
surface access would be less costly. The operational costs would be about the
same as for individually pumped dewatering wells, assuming the same well-
effectiveness and water qualities as before. It is readily apparent that
this type of system would not be cost-effective for control of AMD at the
Lancashire No. 20 mine.
A dewatering system that appears to be cost-effective would be a system
of inclined drainage holes 6.4 cm (2.5 in.) in diameter, drilled with conven-
tional mine equipment. The drilling machine would be mounted on a mobile
column platform and would use 0.19 m^/s (400 cfm) of 690 x 10J Pa (100 psi)
air to drill with. Capital costs were estimated on a footage basis and are
summarized in Table 27. The drain holes would be drilled into the mine roof
from entries around the periphery of the mine and along longwall panels at
angles that would intersect fractures and drain by gravity into a header pipe.
115
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TABLE 27. CAPITAL COSTS FOR INCLINED DRAINAGE HOLES
Item
Equipment
Labor
Expendable Materials
Air, oil, etc.
Bits
Cost (per meter)
$0.85
2.16
2.95
1.02
0.26
$7.24
Cost (per foot)
$0.26
0.66
0.90
0.31
0.08.
$2.21
From the data available, there is no way of predicting the flow from
these holes, but a break-even yield can be calculated for capital costs in
terms of flow per 30 m (100 ft) of drain hole. It can be assumed that the
collection of water from these holes into header pipes and eventually into
a sump or sumps for pumpage to the surface would cost about the same as cur-
rent underground water collection. Therefore, the cost reduction would be
in treatment costs. From Table 17 the capital costs for water treatment are
$430,000. At a cost of $7.24 per meter ($2.21 per foot), 59 300 m (194,600 ft)
of drain hole can be drilled for about the same cost. The yield per 30 m
(100 ft) of drain hole to handle 11 700 mj/d (3.1 mgd) or 135 Us (2,153 gpm)
would be 135 7 1946 or 0.07 l/s. This yield per 30 m (100 feet) of drain hole
appears to be'a reasonable yield to attain and it could easily be exceeded if
the drain holes are positioned to intercept fracture zones or major joint
sets. If the cost of well installation and operation only breaks even with
capital costs for water treatment, there would still be a savings in operating
costs for water treatment. This system is flexible and discharges can be made
to the surface through boreholes where convenient. Actually this system could
result in additional savings in water collection costs since it would require
less piping, and other mine drainage that is good quality could be combined
with the drain hole water and discharged without treatment. Also, this system
is not sensitive to mine depth because the cost of drainage holes would be
the same for all depths. Therefore, this system compares favorably with the
existing system and appears to be a cost-effective approach, especially if
other benefits can be realized.
116
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REFERENCES
1. R. S. Klingensmith, Pennsylvania's Problems and Progress in Mine Drainage
Control, paper presented at 38th Annual Conference, Water Pollution Control
Federation, Mine Drainage and Treatment Plant. AIME Soc. of Mining
Engineers Preprint No. 67F35, 1967.
2. See for example: R. R. Parizek, Prevention of Coal Mine Drainage Formation
by Well Dewatering, Pennsylvania State University Special Research Report
SR-82, 1971, 73 pp; H. L. Lovell and J. W. Gunnett, Hydrogeological Influ-
ences in Preventive Control of Mine Drainage from Deep Coal Mining,
Pennsylvania State University Special Research Report SR-100, 1974, 96 pp;
F. T. Caruccio and R. R. Parizek, An Evaluation of Factors Influencing
Acid Mine Drainage Production from Various Strata of the Allegheny Group
and the Ground Water Interactions in Selected Areas of Western
Pennsylvania, Pennsylvania State University Special Research Report SR-65,
1967, 213 pp.
3. H. J. Gluskoter, Composition of Ground Water Associated with Coal in
Illinois and Indiana, Economic Geology, vol. 60, 1965, pp. 614-620.
4. R. L. Loofbourow and R. L. Brittain, Dewatering through Wells before Mine
Development, Mining Congress Journal, July 1964, pp.43-50.
5. M. R. Campbell, F. G. Clapp, and C. Butts, Description of the Barnesboro
and Parton Quadrangles, USGS Geologic Atlas, Folio 189, 13pp.'
6. N. K. Flint, Geology and Mineral Resources of Southern Somerset County,
Pennsylvania - County Report C56A, Pennsylvania Geological; Survey,
Harrisburg, Pa., 1965, 267 pp.
7. W. E. Edmunds, Revised Lithostratigraphic Nomenclature of the Pottsville
and Allegheny Groups (Pennsylvania), Clearfield County, Pa., Pennsylvania
Geol. Survey, Fourth Series Inf. Circ. No. 61, 1969.
8. N. K. Flint, Geology and Mineral Resources of Southern Somerset County,
Pennsylvania - County Report C56A, Pennsylvania Geological Survey,
Harrisburg, Pa., 1965, 267 pp.
9. C. V. Theis and C. S. Conover, Chart for Determination of the Percentage
of Pumped Water Being Diverted from a Stream or Drain, Shortcuts and
Special Problems in Aquifer Tests (compiler R. Benthall), USGS Water
Supply Paper 1545-C, 1963, pp. 106-109.
117
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10. C. V. Theis and C. S. Conover, Chart for Determination of the Percentage
of Pumped Water Being Diverted from a Stream or Drain, Shortcuts and
Special Problems in Aquifer Tests (compiler R. Benthall), USGS Water
Supply Paper 1545-C, 1963, pp. 106-109.
11. R. R. Parizek, Prevention of Coal Mine Drainage Formation by Well De-
watering, Pennsylvania State University Special Research Report SR-82,
1971, 73 pp. See also Bibliography passim.
12. Skelly and Loy, Inc., Process, Procedures, and Methods to Control Pol-
lution from Mining Activities, U.S. Environmental Protection Agency,
EPA 430/9-73-001, 1973.
13.' Monterey Coal Company (Houston, Texas), 1976, Survey of Accounting
Principles, Procedures, and Practices in the Coal Industry, Proc. of
Second Symposium on Coal Management Techniques, held at Louisville, Ky.,
on October 19-21, 1976.
14. H. B. Charmbury, D. R. Maneval, and C. Girard, Operation Yellowboy -
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TECHNICAL REPORT DATA
(Please read Instructions on the reverse before completing)
1, REPORT NO.
EPA-600/7-79-124
3. RECIPIENT'S ACCESSIOI*NO.
4. TITLE AND SUBTITLE
DEWATERING ACTIVE UNDERGROUND COAL MINES:
TECHNICAL ASPECTS AND COST-EFFECTIVENESS
5. REPORT DATE
July 1979 (issuing date)
6. PERFORMING ORGANIZATION CODE
7. AUTHOR(S)
8. PERFORMING ORGANIZATION REPORT NO.
9. PERFORMING ORGANIZATION NAME AND ADDRESS
W.A. Wahler and Associates
Palo Alto, California 94303
10. PROGRAM ELEMENT NO.
EHE 623
11. CONTRACT/GRANT NO.
68-03-2366
12. SPONSORING AGENCY NAME AND ADDRESS
Industrial Environmental Research Lab-Cincinnati, Ohio
Office of Research and Development
U.S. Environmental Protection Agency
Cincinnati, Ohio 45268
13. TYPE OF REPORT AND PERIOD COVERED
Final 12/75 •- 6/78
14. SPONSORING AGENCY CODE
ET>A/600/12
16. SUPPLEMENTARY NOTES
16. ABSTRACT
This study evaluated the cost-effectiveness of dewatering an active underground
coal mine as an alternative or supplement to treating acid mine drainage. Through
a combination of research, personal contacts and site visits to selected wet mines
in the northern part of the eastern bituminous coal mining region, the Barnes and
Tucker Company's Lancashire No. 20 mine, located in western Pennsylvania, was
selected as a suitable site for a pilot-scale dewatering operation. A dewatering
program was formulated and base-line data collection was performed in conjunction
with exploration of hydrogeologic conditions and well construction. Dewatering
was then performed for a total of 25 days to provide the basis for the technical
and cost analysis.
17.
KEY WORDS AND DOCUMENT ANALYSIS
DESCRIPTORS
b. IDENTIFIERS/OPEN ENDED TERMS C. COSATI Field/Group
Underground Mining
Surface Waters
Ground Waters
Water Quality
Economics
Coal Mining
Acid Mine Drainage
Dewatering
Active Coal Mines
43B
48A
48G
91A
18. DISTRIBUTION STATEMENT
Release to Public
19. SECURITY CLASS (ThisReport}
Unclassified
21. NO. OF PAGES
138
20. SECURITY CLASS (Thispage)
Unclassified
22. PRICE
EPA Perm 2220-1 (9-73)
124
•&• U. S. GOVERNMENT PRINTING OFFICE: 1979 — 657-060/5355
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