EPA-670/2-75-047
June 1975
Environmental Protection Technology Series
p«__^^__^M«_
UP-DIP VERSUS DOWN-DIP MINING
An Evaluation
National Environmental Research Center
Office of Research and Development
U.S. Environmental Protection Agency
Cincinnati, Ohio 45268
-------
EPA-670/2-75-047
June 1975
UP-DIP VERSUS DOWN-DIP MINING
AN EVALUATION
by
John W. Mentz and Jamison B. Wang
Skelly and Loy, Engineers-Consultants
Harrisburg, Pennsylvania 1711O
Program Element No. 1BB040
Project No. 68-O1-0465
Project Officer
Eugene F. Harris
Mining Pollution Control Branch (Cincinnati, Ohio)
Industrial Waste Treatment Research Laboratory
Edison, New Jersey O8817
NATIONAL ENVIRONMENTAL RESEARCH CENTER
OFFICE OF RESEARCH AND DEVELOPMENT
U.S. ENVIRONMENTAL PROTECTION AGENCY
CINCINNATI, OHIO 45268
-------
REVIEW NOTICE
The National Environmental Research Center--Cincinnati
has reviewed this report and approved its 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 recommen-
dation for use.
11
-------
FOREWORD
Man and his environment must be protected from the adverse
effects of pesticides, radiation, noise and other forms of pollution,
and the unwise management of solid waste. Efforts to protect the
environment require a focus that recognizes the interplay between
the components of our physical environmentair, water, and land.
The National Environmental Research Centers provide this multi-
disciplinary focus through programs engaged in
studies on the effects of environmental contaminants on
man and the biosphere, and
a search for ways to prevent contamination and to recycle
valuable resources.
This report presents detailed results of a feasibility study of
down-dip mining, a technique that appears to offer an alternative to
sealing or permanant treatment of polluted effluents from under-
ground coal mines after abandonment. Of special importance in the
study were evaluation of pollution reduction effectiveness of the
technique and the National water quality impacts of its widespread
employment.
A. W. Breidenbach, Ph.D.
Director
National Environmental
Research Center, Cincinnati
-------
ABSTRACT
The report presents detailed results of a feasibility study of down-
dip mining, a technique that appears to offer an alternative to sealing
or permanent treatment of polluted effluents from coal mines after
abandonment. The project included an evaluation of pollution reduc-
tion effectiveness, economic and engineering limitations, costs in
varying situations, health and safety aspects, and National economic
and water quality impacts of utilization of this technique.
Project efforts included location and evaluation of a pair of aban-
doned underground mines - one developed to rise, one developed to
dtp - closely similar in all other aspects. The principal goal of this
portion of the project was confirmation of the theory that discharge
water quality in down-dip mines is substantially better than that in
up-dip mines. An active mine with units operating up-dip and down-
dip was also evaluated to ascertain major advantages and disadvan-
tages of each mode of operation. Key factors which could vary with
different mining procedures were pumping and coal haulage. How-
ever, haulage cost differentials were found to be extremely small,
and pumping costs, which are related to highly variable local infil-
tration rates, could not be predicted. Health and safety and National
water quality and economic impacts were also evaluated. Findings of
each phase of study are presented here.
This report was submitted in fulfillment of Grant Project Number 68-
01-0465 by Skelly and Ley, Engineers-Consultants, under the spon-
sorship of the Environmental Protection Agency. Work was com-
pleted as of January 31, 1975.
iv
-------
TABLE OF CONTENTS
Page
ACKNOWLEDGMENTS viii
I CONC LUSIONS 1
II RECOMMENDATIONS 3
III INTRODUCTION 4
IV ABANDONED MINE SITE EVALUATIONS 10
Selection Criteria 10
Mine Locations 10
Geology 11
Mining History 14
Shoff Mine 14
Yorkshire No. 1 Mine 17
Mine Drainage and Water Quality 18
Shoff Mine 24
Yorkshire No. 1 Mine 24
V ACTIVE MINE SITE EVALUATION 28
Mine Location 28
Geology 28
Mine Development 30
Mining and Production 31
Mining Technique 31
Production 33
Mine Drainage and Water Quality 36
VI ANALYSIS OF DOWN-DIP MINING TECHNIQUES . . 41
Water Quality 41
Production 42
Economics 48
Health and Safety 60
National Impact 61
Feasibility and Applicability 64
VII REFERENCES 70
VIII DEFINITION OF TERMS 71
v
-------
LIST OF FIGURES
Figure Page
1 Location Map 9
2 Yorkshire No. 1 Mine and Shoff Mine -
Location and Geology 12
3 Stratigraphic Column of Surface Rock 13
4 Yorkshire No. 1 Mine and Shoff Mine -
Mine Development 15
5 Stott No. 1 Mine 29
6 Stott No. 1 Mine - Raw Coal Production Before
and After Health and Safety Act of 1969 43
7 Stott No. 1 Mine - Raw Coal Productivity 44
8 Stott No. 1 Mine - Monthly Unit Production
Rates - 1973 46
9 Percent Increases in Cost Per kkg for 61 cm.
(24 in.) Belt Coal Haulage Up-grade Versus
Down-grade 55
1O Percent Increases in Cost Per kkg for 76 cm.
(3O in.) Belt Coal Haulage Up-grade Versus
Down-grade 56
11 Normal Down-dip Mining Procedure . 68
12 Alternate Down-dip Mining Procedure 69
VI
-------
LIST OF TABLES
Table Page
1 Historical Water Quality Data - Shoff and
Yorkshire No. 1 Mines 20
2 Abandoned Mine Water Quality Data - Shoff
and Yorkshire No. 1 Mines 21
3 Summary - Abandoned Mine Site Characteristics 27
4 Stott No. 1 Mine - 1973 Production Record 34
5 Stott No. 1 Mine - 1967-68 Production Record 38
6 Stott No. 1 Mine - Power Consumption Breakdown .... 50
7 Conveyor Belt Operation Costs - Up-grade Versus
Down-grade Haulage 52
8 Percentage Cost Increase in Up-grade over Down-
grade Conveyor Haulage 54
9 Conversions Table-English to Metric 73
-------
ACKNOWLEDGMENTS
Cooperation and special assistance for development of information
from the Stott No. 1 Mine was provided by Mrs. Frances C. Stott,
President of Lady Jane Collieries, Incorporated.
Messrs. Norman Wetzel, Charles Merritt and Raymond Alvertro
provided valuable data and technical support during the active mine
site investigative portion of this study.
Information from numerous Federal and State agencies, referenced
in this report, is gratefully acknowledged. The following persons
also provided valuable input to this project:
Luke Amos, Chief Engineer of L. B. Smith, Incorporated,
Don Caldwell, Division of Water Resources, West Virginia
Department of Natural Resources,
Ralph Dean, State Deep Mine Inspector, Ohio Department of Natural
Resources
Richard D. Thompson, Division of Industrial Waste and Erosion
Regulation, Pennsylvania Department of
Environmental Resources.
Principal investigators in this study were Jamison B. Warg and
John W. Mentz.
VI11
-------
SECTION I
CONCLUSIONS
Based upon the findings of this study, the following conclusions have
been made:
Acidity concentrations in the main discharge from the abandoned
mine developed to the rise were consistently in excess of 2200
mg/l, while those of the discharge from the mine developed to the
dip were generally less than 10O mg/l. Iron, sulfate, manganese
and aluminum concentrations were also significantly higher in dis-
charges from the up-dip mine.
Since the abandoned mines evaluated in this study were similar in
all other respects, the primary factor controlling water quality
can only be the direction of mine development.
Employment of down-dip mining techniques will cause no signifi-
cant water quality improvements on a National scale, since re-
cently proposed Federal effluent limitations guidelines for the
coal industry will require all active mining operations to maintain
acceptable water quality.
Pollution formation reductions attributable to down-dip mining will
be of major economic importance to mine operators who, as a
result of proposed effluent guidelines, are faced with the prospect
of perpetual treatment of discharges from their mines after
closure.
Implementation of down-dip mining in major acid-producing coal
regions could have significant regional and National impact on
treatment costs, which are tied directly to production costs.
The minimal concentrations of acid mine drainage observed within
the down-dip mine are attributed to local site conditions, and do
not reflect any deficiency in the pollution reduction potentials of
the down-dip mining technique.
When all pertinent mining factors are considered, down-dip mining
at the active site was no more or less productive than up-dip
mining.
-------
Numerous physical factors (including geologic and hydrologic con-
ditions), which are only marginally related to mine development or
mining technique, can have combined effects that are sufficient to
obscure any advantage or disadvantage of mining to the dip.
Substantial declines in both total production and productivity per
man-shift occurred at the active minesite after passage of the
Health and Safety Act of 1969.
In the active mine evaluated, coal haulage is by far the largest
single power cost.
For a constant length belt segment operating at optimum speed for
coal haulage, cost differentials for haulage up-grade (as in a down-
dip developing mine) over haulage down-grade (as in an up-dip
mine) at any realistic grade angle is generally less than one half
cent per metric ton - a relatively minimal economic impact.
Despite extremely low conveyor belt haulage unit cost, percentage
cost increases incurred by haulage up-grade rather than down-
grade range from a few to several hundred percent.
Pumping costs, which vary greatly according to geology, hydrol-
ogy, techniques employed, and mining plan, may be substantially
increased by down-dip mine development; but they will probably
not reach the point of adversely affecting production economics.
Effective sealing of mines in steeply dipping seams can only be
assumed where mines have been developed to the dip and the lower
outcrop or barrier integrity has been maintained.
Since down-dip mining operations must remain in compliance with
all facets of the Health and Safety Act of 1969, there are no ad-
verse health and safety considerations which could affect utilization
of this technique.
Feasibility and applicability of down-dip mining techniques have al-
ready been established in numerous applications under various
mining conditions.
Alternative mining procedures are available to counter any in-
creased water handling attributable to down-dip mining. The cap-
ital investment required to implement these techniques may, how-
ever, be higher than the benefits realized.
-------
SECTION II
RECOMMENDATIONS
Down-dip mining techniques should be implemented wherever they
may have significant beneficial pollution reduction effects on dis-
charge water quality.
An in-depth evaluation of the alternate down-dip mining procedures
proposed in this report should be conducted to determine their ap-
plicability, feasibility, productivity, and economics.
Detailed analysis of down-dip barrier pillar and coal outcrop
strength characteristics, permeability, variables controlling these
factors, and procedures for estimating required thicknesses would
prove beneficial in maximizing coal recovery while minimizing
chances for outcrop or barrier failure.
-------
SECTION III
INTRODUCTION
The fact that at least two thirds of all coal related acid mine drainage
discharges from underground mines has been well documented by
numerous studies. It has also been established that a very high per-
centage (75 - 90%) of this underground mine drainage pollution ema-
nates from abandoned mine sites. Reasons for the high level of pol-
lution from abandoned underground mines are largely related to
mining methods utilized in the past. Historically, the oldest undei
ground coal mines were generally drift mines; that is, they entered
the coal seam where it intersected or outcropped on the land surface.
It was also common to open the drift on a low side of the coal seam
(the down-dip side), and mine toward the rise of the coal. This per-
mitted easy haulage of loaded coal carts from the working coal face
down-dip to the mine entry. In addition, as the mine was developed,
most water infiltrating into the workings would drain by gravity,
thereby minimizing the amount of pumping required to keep active
areas dry.
Groundwater entering the mine workings frequently becomes polluted,
and up-dip mining practices permit this mine drainage to flow freely
from the mine both during active operations and after abandonment.
Typically, mine atmospheres exhibit humidity near the saturation
point. Thus, regardless of other sources of water, there is always
moisture condensing on mine walls, roof and floor. This moisture,
combined with oxygen in the mine atmosphere and pyritic materials
in the coal, floor material, or roof rock, forms acid salts. The
acid "weeps" down walls of the workings as condensation continues
and newly condensed moisture replaces acid-bearing moisture. When
acid reaches the floor of the workings, it is transported to another
portion of the mine, or from the mine, by flowing water. This
flowing water, usually the result of groundwater infiltration, may be
steady or intermittent in nature depending on climatic conditions. In
either case, if the mine has been developed to the rise and drains by
gravity, acid drainage will eventually flow from the workings and
subsequently degrade water quality of receiving streams.
Oxidation of pyrite and formation of acid mine drainage are not
problems that decline rapidly in severity with time. Acid formation
and discharge may continue for over a hundred years after closure
-------
as fresh pyritic materials are constantly exposed by minor roof
collapses. In addition, groundwater infiltration and condensation of
moisture from the mine atmosphere are constant sources of water.
Water quality of discharges from acid-producing abandoned mines
will improve only after the mine roof has completely collapsed into
the workings, and pyritic materials have been leached out or sealed
off from further oxidation. Thus, acid formation is a long-term
continuing process and must be dealt with in a manner that will ef-
fectively keep oxygen from entering the mine.
Pollution emanating from active underground mines has continually
decreased with increased effectiveness of water quality controls. In
fact, active mines should soon cease to be a source of pollution when
presently available control mechanisms of state and Federal govern-
ment are fully employed. Most active mines are currently required
to provide some form of treatment to contaminated drainage, and
pollution from active underground mines is not considered a major
problem today. However, upon closure of an underground mine, the
question arises as to responsibilities of former mine operators in
maintaining water quality, and duration of those responsibilities.
The United States Environmental Protection Agency's effluent limi-
tations guidelines for the coal industry, which could be enacted as
early as July 1, 1975, would place permanent responsibility for
maintenance of acceptable water quality in abandoned mine discharges
upon the former operator.
Current technology for maintenance of acceptable water quality from
underground mines is limited to mine sealing and treatment. Many
mines developed to the rise and currently active can never be effec-
tively sealed, because of tremendous hydraulic heads that would
develop as the mines flooded. Heads in excess of 6 or 7 meters
(20 to 23 feet) cannot be safely withheld, either by mine seals or
outcrop barriers (which are often weakened by adjacent surface or
underground mining). In such situations, constant seepage can occur
through the seal or the adjacent coal outcrop barrier. In addition,
there is always the possibility of failure of a seal or adjacent outcrop
barrier, which could, depending upon the volume of water contained,
have devastating effects. Discharges from mines located in acid-
producing coal seams and exhibiting such deterrents to sealing will
require constant treatment - an extremely costly solution to the
mine drainage problem.
-------
This study was initiated to investigate a specific mining technique
which appears to offer an alternative to mine sealing or permanent
treatment of effluents from abandoned coal mines. This method,
referred to in this report as down-dip mining or mining to dip, is
not really new, since it has been employed on a limited scale
throughout the coal fields for many years.
The principle of down-dip mining involves entry to the coal seam at
the highest point of mine development and mining toward lower coal
elevations. In mines so developed, both coal and mine drainage must
be removed from mine workings against the force of gravity. Until
recent years, this technique was not utilized extensively since it
often increased coal production costs substantially. However, im-
provements in mining, pumping, and haulage technology have greatly
reduced cost differentials to the point where it is actually a viable
and competitive alternative mining method.
In theory, one of the principal benefits of mining to dip is not actually
realized until after mine abandonment. All water entering the work-
ings drains down-dip, but there is no discharge point in the low side
of the mine. Therefore, when a mine is abandoned and pumping
ceases, the workings inundate naturally. Flooding isolates pyritic
materials in the coal, overburden, and mine floor. Thus, pyritic
oxidation virtually ceases and acid mine drainage production almost
halts. When the mine has completely filled and begins to discharge
from the entry on the up-dip side, water may still be acid in nature,
until the previously formed pollutants drain from the workings. How-
ever, in a relatively short time, within a few years, discharge
quality will substantially improve because no new acid is being
formed in the workings. In this manner, mining to dip can greatly
reduce costs of maintaining environmental quality after abandonment.
Specific goals of this evaluation of down-dip mining were cited in the
project proposal:
1) Evaluate actual effectiveness of the technique by studying pollution
formation within similar mines developed to rise and to dip.
2) Develop applicability of the technique and its economic and en-
gineering limitations.
3) Determine cost variations that could be incurred by employing the
technique in various situations.
-------
4) Evaluate health and safety aspects of the technique.
5) Evaluate National water quality and economic impact of employing
this technique.
6) Develop design criteria and general feasibility of utilizing this
technique.
7) Generate recommendations for use by regulatory agencies to im-
plement the technique, if feasible.
Water quality improvements attributable to use of down-dip mining
(as opposed to up-dip mining) were determined by comparison and
evaluation of abandoned underground mines. These mines were
selected to be as similar as possible in all aspects, except that one
mine was developed to rise and the other to dip. All available his-
torical water quality data and mine mapping was obtained and closely
scrutinized to ascertain mine conditions and extent of similarity
between the two mines. To evaluate water quality, a periodic sam-
pling and flow measuring program was initiated, with sampling
stations at every point of discharge from both mines.
Initial project plans called for analysis of two active mines, one
operating to rise and one operating to dip, with similar characteris-
tics. Numerous contacts were made in an effort to locate a pair of
suitable mines. State and Federal regulatory agencies concerned
with underground coal mining and State mine inspectors from several
regions were interviewed. Background knowledge and experience
gained during completion of two nationwide studies of the coal in-
dustry were also utilized. Contacts with coal companies or firms
operating captive coal mines included Consolidation Coal Company,
Jones and Laughlin Steel Corporation, Bethlehem Mines Corporation,
Peabody Coal Company, Southern Ohio Coal Company, Bradford
Coal Company, and Lady Jane Collieries, Incorporated.
This search for active mining operations suitable for investigation
revealed a number of important points. First, a cooperative pair of
mines sufficiently similar to warrant further consideration could not
be located. In many cases, mines are so large that they concurrently
include both up-dip and down-dip development, and cannot be accu-
rately evaluated precisely as up- or down-dip operations. On the
other hand, mines small enough to be developed exclusively to rise or
-------
dip are uncommon and generally do not maintain sufficiently detailed,
usable data on production or economics.
Based on these preliminary findings, consideration was given to
evaluation of a single large mine in which different sections being
developed to rise or to dip could be isolated. Production and eco-
nomics in a large mine could be evaluated for either single production
units or panel segments of the mine. Several coal producers con-
tacted simply had no operations that adequately fit requirements of
this study. Some did not have sufficiently detailed production and
economic data breakdowns to permit analysis of different mine
sections. Other mine operators were unwilling to divulge information
they considered to be confidential, including specific production rates
and detailed cost breakdowns. Some tentatively agreed to cooperate
early in the study only to decline after discovering the level of detail
required in data being sought. A recent deluge of environmentally
oriented studies and surveys, many quite demanding of the coal in-
dustry* also made many companies hesitant to cooperate in yet an-
other evaluation, regardless of its purpose. In fact, of the com-
panies contacted, only Lady Jane Collieries, Incorporated, agreed to
cooperate in this study. Fortunately, while not totally ideal, Lady
Jane met most of the prerequisites established for a meaningful eval-
uation. Company representatives provided valuable production infor-
mation and insights into economics of mining coal. Based on this
information, other goals of the study were achieved and recommen-
dations were developed. Findings of this study and all pertinent data
are presented in this report.
8
-------
tnll.ld/V
£ *LADY JANE COLLIERIES
ui
CLEARFIELD
CENTRE
LPhilipsburg
YORKSHIRE V
US.322
State College
MINE
Coalport
BlandburgrTyronc
A Itoona
lohnstown
SOMERSET
Chamber iburq
LOCATION MAP
Seal* m MM*t
Figur* 1
-------
SECTION IV
ABANDONED MINE SITE
EVALUATIONS
SELECTION CRITERIA
The primary basis for selection of the two abandoned mines studied
in this project was their great degree of similarity. Such similari-
ties were mandatory for this project in order to permit the most
dependable evaluation of relative water quality. Both mines are
similar in all of the following parameters: coal seam mined, coal
quality (particularly sulfur content), approximate mine size, mining
methods utilized, time period of operation, availability of mine
history and msfpping, geologic controls, hydrologic setting, topo-
graphic regime, and measureable discharges.
Abandoned mines selected for this in-depth study were the Shoff and
Yorkshire No. 1 Mines in Clearfield County, Pennsylvania. The
Shoff Mine was worked to rise, the most common mining procedure
in past years, while the Yorkshire No. 1 Mine, hereafter referred to
simply as the Yorkshire Mine, was operated to dip. Since abandoned
mines worked to the rise are almost always greater sources of acid
mine drainage, the Shoff Mine serves as a control to determine if the
water quality of the down-dip Yorkshire Mine shows a difference in
pollution load. Table 3, page 27, shows summary information.
MINE LOCATIONS
The abandoned workings of the Shoff Mine and the Yorkshire Mine are
located in Clearfield County, Pennsylvania, just west of Madera.
Both mines are in Bigler Township, and lie on opposite banks of
Clearfield Creek. Both Shoff and Yorkshire Mines have major dis-
charges which enter Clearfield Creek just upstream from the mouth
of Muddy Run, a tributary highly polluted by acid mine drainage.
The Yorkshire Mine, which underlies approximately 220 hectares
(544 acres), has a pair of drift entries just south of Clearfield Creek.
-------
There are also two slope entries one kilometer (two-thirds of a mile)
southeast of the drifts, adjacent to the mouth of Banian Run. The
mine extends approximately 3.2 kilometers (two miles) south and
southwest from the drifts, ranges in width from 30O to 1830 meters
(1OOO to 6000 feet), and underlies several small hills and portions of
Banian Run.
The Shoff Mine workings lie only about 61O meters (200O feet) north
of the Yorkshire drift entries on the north bank of Clearfield Creek.
Slightly smaller than Yorkshire, the Shoff. Mine occupies 170 hec-
tares (428 acres) beneath a single, large hill. The mine extends 1980
meters (65OO feet) north from Clearfield Creek, and ranges from 490
to 1650 meters (6OO to 5400 feet) in width. Thirteen drift entries to
the mine complex are located along its southern and eastern edges.
Locations, general configurations and entries of the Shoff and York-
shire Mines are shown in Figure 2.
GEOLOGY
The Shoff and Yorkshire Mines were both developed in the highly
pyritic Clarion coal, or "A" seam, which outcrops about 12 meters
above stream level on both the north and south banks of Clearfield
Creek. Stratigraphically the lowest Allegheny Group coal, the
Clarion is underlain by sandstones comprising the upper portion of
the Pottsville Group. All of the other major Allegheny Group coal
seams - the Lower Kittanning "B" coal, the Middle Kittanning "C"
coal, the Upper Kittanning "C - Prime" coal, the Lower Freeport
"D" coal, and the Upper Freeport "E" coal - outcrop on the hillsides
above both mine workings. All but the Upper Kittanning coal are at
least locally strippable in this area, but additional deep mining was
limited to some small Lower Freeport workings on the hilltop above
the Shoff Mine. Figure 3 is a stratigraphic column showing relative
positions and types of strata found in the study area.
11
-------
YORKSHIRE
SHOFF
Figure 2
12
-------
SYSTEM
CO
O
hi
b.
Z
FORMATION
Motioning
Frtcporl
Opplr
K.rtannmg
MlddK
Kiltanning
Lo««r
KillOnnine
Clarion
Uorctr
Pocono
COAL
MEMBER
Mohonina
Upp*r Fmport
Untr Friiport
Lower Frttport
Upp«r
Kittanning
MiddM
Kittanning
LOW
Killonninfl
Clarion
Morctr
SECTION
RHMM
*^fc~ "
^
E?=I==?=-"=^i."^
B^5s^Sera^«
s^=^r=s=r==i==£:
CHARACTERISTICS OF
COAL MEMBER
Aviroflt thtcHAMt" 1 foot. Thin ond trtofolly
diuonliflu«ut.Stripmin*d locglly.
Eitmiwily ttrip mlnid *Hfi wnw dt>p mining.
Split into i BinchM Kith o SO foot ttparotiofl.
Both art laterally conllnuoui with avirogi
IWcknul of lOlnchn and or* t»t«nt,»«l»
trip mlflod and tfoip mined.
Avtraga thjehnoit1 24inchti.ljotfrally dlt-
coilinuoui but locally itrip and daip mlnad.
Avlragi thickniu -36inthM Folr pirtOtancl.
Motlli itrio minad «ith lomi dMD mlnlna.
Aviragi tnicknen M Incrwt. Including a 10
inch tlwlo parting. Lattrolly contlnMui and >-
Avaragi mckniii -t2lnch«. Lanroliy
coatinuota. Ottp nlntd ond itrlp mlnad.
Vary thin and loNrally diiconllnuoui. Locally
ttrip nin«d.
GENERAL STRATIGRAPHY
OF GROUP
Errallc cyclic uquirott of undttoni,
IIIHIc.nl, mow ond thin coalt, Lo*«r 23S
Kit of group.
Veriabl* cyclic Mquinco of clay.cloyltoni,
ond ninoablo cooli. Avarogo thieknni
Z99 («lt
Poorly dwaiopod cyclic tiqutnca of fhoiot,
tilttton«i,tand*tonos, ond thin eoolt.
Thickntn -SO foot.
Flni-graintd to conglomorotlc tonditont
up to JSO f»l thick.
VfRIICAL SCAIE
so g ipo F«I
STRATIGRAPHIC COLUMN OF SURFACE ROCK
Figure 3
13
-------
Geology of this portion of central Pennsylvania has been formally
mapped by Mr. William Edmunds, Chief Coal Geologist for the
Pennsylvania State Geological Survey. Mr. Edmund's mapping
shows the Shoff and Yorkshire Mines are structurally situated on the
southeast limb of the northeast-southwest trending Laurel Hill Anti-
cline. The Shoff workings are centered only 1.6 kilometers (one
mile) southeast of the anticlinal axis, and the Yorkshire workings are
only 0.8 kilometer (one-half mile) farther away. As a result, strata
in the vicinity of both mines dip south or southeast at about one de-
gree, as shown by the structure contours in the Figure 2 mapping.
There are numerous large faults just north of the abandoned mines,
along the anticlinal axis. However, no major faults are known to
exist in the immediate vicinity of either mine.
MINING HISTORY
Shoff Mine
The Clarion coal seam at the Shoff Mine site was mined at different
times by both underground and surface techniques between the late
19th century and the early 193O's. During the 194O's, approximately
460 meters (1500 feet) of "A" coal outcrop adjacent to the workings,
and 610 meters (2000 feet) of the overlying "B" coal outcrop east and
southeast of the mine were stripped. In addition, several small "D"
and "E" coal strip cuts were made above the Shoff Mine complex.
Some of this stripping has since been backfilled, regraded and re-
vegetated, and now supports well established plant growth. There
are currently no active mines in the vicinity of the Shoff Mine. All
mining information indicates the Shoff Mine was developed to the rise
to facilitate gravity drainage from workings and to take advantage of
rail facilities located adjacent to the creek. Although most entries
have been destroyed by road building activities along Route 53, re-
view of available mine maps show that there were 17 drift entries
driven from the south and east. The southernmost drifts were pri-
marily utilized for haulage of coal from the mine to the tipple and
rail sidings along Clearfield Creek. Other entries provided access
for personnel and supplies. A generalized view of the workings and
mine development of the Shoff Mine complex is shown in Figure 4.
14
-------
.
n
TWP.
BIGLER
\
N ->
INUNDATION
v
YORKSHIRE NO. 1 MINE &
SHOFF MINE
Sampling Point
Inundated Area
MINE DEVELOPMENT
1000 0 2000 Fee i
Structure Contours H60 £;
15
-------
The Shoff Mine consists of several blocks of interconnected workings
which were apparently first mined independently from one another.
The southernmost block, originally known as the Gatehouse Mine,
measures approximately 460 by 550 meters (15OO by 18OO feet),
with access through six drift entries. The extreme southern drift in
the Shoff complex was the original point of entry to this block of
workings. From this point, a main heading was driven 460 meters
(1500 feet) to the northeast, roughly paralleling the coal outcrop in
that area. Five headings were then driven to the left, or northwest,
from that main heading, permitting extensive coal extraction.
An interconnected northeastern block of coal was also worked inde-
pendently by the Greenwood Mine. There are three drift entries to
the Greenwood Mine workings, but no discharge has been observed at
any of these. It is suspected the workings are partially inundated and
draining to the northern end of the Shoff Mine.
Most of the Shoff Mine is comprised of the remaining central block of
coal. The mine's main heading, which is now evident only as the
primary drainage discharge point (DD-1) extends approximately 1160
meters (3800 feet) northwest, slightly west of the maximum dip.
From this main heading, four left headings were driven down-dip
southwest. Beyond the workings of the previously mentioned Gate-
house Mine, about 550 meters (18OO feet) in from the drift entry, the
first of these headings bore to the left. After breaking through the
coal outcrop, the first of these left headings was utilized as an entry-
way. This was the westernmost drift in the Shoff Mine. The barrier
between the ends of the three remaining down-dip headings and the
coal outcrop apparently ranges from 15 to 15O meters (50 to 5OO feet).
Large barrier size here probably reflects a property or mineral
rights ownership boundary, since such blocks of coal are not usually
left unmined.
Although these 2nd, 3rd and 4th left headings were driven toward the
dip, they did not break through to the surface, and it is not likely they
are currently inundated. Rooms driven from these headings inter-
connect with one another, allowing drainage southwest to the first left
heading. This heading is presently a significant source of acid mine
drainage.
Seven right headings were driven to the rise from the Shoff Mine's
main heading. These northeast-trending headings have a total com-
16
-------
bined length of about 6.5 kilometers (4 miles). Due to their orienta-
tion, they are probably not flooded. These up-dip headings were
driven on 150 meter (500 foot) centers, while rooms or panels were
driven on 15 meter (5O foot) centers and range in length from 30 to
12O meters (100 to 40O feet). Maximum recovery, or full pillar
extraction, was practiced in the southern portions of the mine. The
lowest recovery rate, about 35%, occurred beneath a maximum of
90 meters (300 feet) cover near the center of the mine. Shoff head-
ings and their associated workings are extensively interconnected,
and combined are the source of mine drainage emanating from the
main entry.
Yorkshire No. 1 Mine
General mine development of the Yorkshire No. 1 Mine is illustrated
in Figure 4, Page 15. Operation of this mine occurred during ap-
proximately the same time period as the Shoff Mine. Portions of
mine mapping for this complex were dated by the original mining
engineers, and indicate that mining began during or shortly before
World War I. During early years of mine development, from about
1916 to 192O, the main heading was driven approximately 1830 meters
(6000 feet) southwest, or down-dip. Eight right headings totalling
roughly 4.4 kilometers (2 3/4 miles) were driven northeast on 90
meter (3OO foot) centers. These right headings were driven to rise
from the main heading with an average grade of 1.5 percent. As in
the Shoff Mine, panels from these headings were driven on 15 meter
(50 foot) centers, and there are indications of only local full pillar
extraction. Available mapping suggests drift and slope entries to the
Yorkshire workings were developed very early in the mine's history.
Coal was transported from the mine through slope entries to a tipple
and rail facilities adjacent to Banian Run. The drift entries were
probably used for personnel and equipment access.
Between 1921 and 1929, the mine's main heading was extended an ad-
ditional 146O meters (4800 feet), but was angled 30 degrees west from
the original heading. Along this new portion of heading, six additional
right headings - the 9th through 14th - with a combined length of 1.9
kilometers (1.2 miles) were driven. In this newer portion of the
mine, the Clarion coal rolls gently, but the general direction of dip
is still to the south.
17
-------
Southeastern portions of the mine were last to be developed, be-
tween 1929 and 1941. Four south headings were advanced down-dtp
off the main heading. The 4th South heading was over eight-tenths
kilometer (one-half mile) in length, and opened a large block of coal.
A second block of coal beneath and southeast of Banian Run was also
mined at this time. Such mining under the stream was limited to
areas in which cover exceeded 21 meters (70 feet), thereby reducing
possibility of excessive groundwater infiltration.
MINE DRAINAGE AND WATER QUALITY
One of the primary goals of this study was determination of effec-
tiveness of mine flooding as a method to reduce or eliminate produc-
tion of acid mine drainage in abandoned underground coal mines.
The general theory behind this pollution reduction technique is dis-
cussed in the introduction to this report. However, the only accurate
way to determine validity of this theory is to evaluate quality of ef-
fluent from a mine which employed down-dip mining as compared to
water quality from an unflooded, up-dip mine's discharge. This was
achieved by monitoring flow and quality of all mine drainage dis-
charges from two selected abandoned mines - the Shoff Mine, un-
flooded with its mine workings above the points of discharge (up-dip),
and the inundated Yorkshire Mine with its development to dip below
the points of discharge. This water quality data, augmented by all
available information on development of the abandoned mines and
local structure of the coal, provided an excellent picture of the mines'
drainage patterns and effects of those patterns on formation of acid
mine drainage.
Historical monitoring data, as well as current sampling results, were
utilized in this study to provide a comprehensive water quality data
base. Historical data was obtained for both Shoff and Yorkshire
Mines from previous sampling programs conducted during studies of
the Clearfield Creek and Muddy Run watersheds. Grab type sam-
pling was commonly used in both programs, and samples were ana-
lyzed for pH, hot acidity, total iron and sulfates. Every Shoff Mine
discharge point was located and sampled during the Clearfield Creek
study. Unfortunately, the main discharge from the Yorkshire Mine
was not sampled in either study, although the small discharge from
tine slope entry was. The purpose of these watershed studies was
18
-------
elimination or reduction of acid emanating from various sources,
including deep mines. Mine drainage with little or no acidity and
relatively high pH, 5.0 or greater, was not monitored consistently
or considered a major source of pollution. The historical water
quality data obtained during those sampling programs is presented
in Table 1.
For this study, monitoring stations were established at all discharge
points for each mine, five for the Shoff Mine and two for the York-
shire Mine. These sampling points are identified in Figure 2, Page
12. Eight semi-monthly sample runs were conducted for collection
of discharge samples and metering flows. All water samples were
analyzed for: pH, acidity, alkalinity, sulfates, ferric iron, ferrous
iron, specific conductance, manganese, aluminum and calcium. Data
obtained during the current sampling program is summarized in
Table 2. Chemical analyses of samples taken early in the study
period were examined closely to evaluate noticeable differences be-
tween field and laboratory pHs. This also raised questions as to the
validity of other analysis data. Inconsistencies were corrected
through comparisons with other quality control analyses made by a
separate laboratory. The laboratory providing the most dependable
work was retained for all further water quality analyses.
A comparison of quality of major discharges from the abandoned
mines provides helpful insights into the hydrogeologic conditions con-
trolling formation of acid in those mines. Acid concentrations in the
Shoff Mine's main discharge were consistently in excess of 2200
mg/l, while those of the discharge from the Yorkshire Mine were
generally less than 1OO mg/l. Similar differences were noted in
other mine drainage indicators. Total iron concentrations ranged
from 6OO to 80O mg/l and sulfates averaged 200O mg/l in the Shoff
Mine, while concentrations of 55 mg/l of iron and 600 mg/l of sulfate
were observed in Yorkshire Mine discharges. Manganese and alumi-
num concentrations were also substantially higher in the Shoff Mine.
Total pollutant loads for the Shoff and Yorkshire Mines were com-
puted combining all discharges from each mine. These combined
loadings, which were presented in Table 3, confirm the historical
water quality data, and are in agreement with water quality analyses
reflected in the main discharge samples.
19
-------
Table 1
HISTORICAL WATER QUALITY DATA
SHOFF AND YORKSHIRE NO. 1 MINES
Discharge
Number
DD-1
DD-2
DD-3
DD-4
DD-5
DD-7
Date
9/14/72
10/80/72
11/O1/72
10/O2/72
11/O1/72
1O/O2/72
10/20/72
11/01/72
9/14/72
10/2O/72
11/01/72
10/02/72
10/20/72
7/30/7O
8/31/7O
9/3O/70
1O/29/7O
1 1/29/70
12/28/70
1/23/71
3/05/71
4/08/71
5/1 1/71
Lab
PH
2.9
2.6
3.1
3.3
2.7
3.4
2.4
2.9
2.8
2.6
3.1
3.6
2.7
3.7
3.7
3.7
3.7
3.9
4.0
4.2
3.9
3.7
3.6
Row
cu.m./min.
(cu.ft./secj^
0.65
(1.45)
0.63
(1.39)
0.24
(0.53)
O.O3
(0.06)
0.09
(O.20)
O.1O
(0.22)
O. 17
(0.38)
O.O9
(O.2O)
O. 1O
(0.22)
O.O3
(O.O6)
O.12
(0.26)
0.03
(O.O6)
O.O8
(0.18)
0.02
(0.04)
O.O7
(0. 15)
O.O3
(0.06)
0.17
(0.38)
0.34
(0.75)
0.78
(1.72)
O.17
(O.38)
1.O5
(2.31)
0.44
(0.97)
O.29
(0.64)
Net Acidity
jrig/l
2870
32OO
2693
1510
158O
2200
1820
1570
1320
1500
1390
7600
34O
62
55
80
1O7
1OO
60
102
66
60
48
kg/day
(Ibs/day)
2737
(6O3O)
2895
(6378)
921
(2O29)
74
(163)
193
(425)
323
(712)
445
(93O)
192
(423)
194
(427)
73
(161)
233
(524)
372
(820)
41
(90)
2
(4)
5
(11)
4
(9)
26
(57)
49
(108)
67
(148)
25
(55)
1OO
(22O)
38
(84)
2O
(44)
Total I ron
mg/l
999
200
93
278
288
338
172
744
376
354
328
173
84
O.7
0.4
0.5
0.3
O.3
.0
.2
O.3
0.1
O.O8
kg/day
(Ibs/day
952
(2O97)
181
(399)
342
(753)
14
(31)
35
(77)
5O
(110)
42
(93)
91
(200)
55
(121)
17
(37)
56
(123)
85
(187)
10
(22)
0.0
O.O
O.O
0.1
(0.22)
O.1
(O.22)
0.0
O.O
0.4
(0.88)
O.1
(0.22)
O.O
Sulfates
mg/l
1970
230O
2500
2460
194O
4OOO
28OO
25OO
1O3O
1450
235O
1330
320
47O
298
47O
461
5O9
461
557
154
336
288
kg/ day
(Ibs/day
1878
(4137)
2O81
(4584)
856
(1886)
120
(264)
237
(522)
587
(1293)
685
(1509)
3O6
(674)
151
(333)
71
(156)
402
(896)
65
(143)
39
(86)
11
(24)
29
(64)
23
(51)
113
(249)
249
(549)
518
(1141)
136
(300)
233
(513)
213
(469)
120
(264)
20
-------
Table 2
ABANDONED MINE WATER QUALITY DATA
SHOFF MINE
SAMPLING STATION DD-I
DATE
7-30-74
8-SO-74
9-OB-74
9-S4-74
10-10-74
10-31-74
11-27-74
12-O2-74
FLOW
Cu m/mjn
0.65
0.20
0.20
O.36
0.09
0.17
0.34
0.19
PH
a
*
-
2.8
_
2.S
2.6
2.5
2.5
2.3
CO
4
_J
2.3
2.6
2.9
2.9
2.7
2.5
2.9
2.7
ACIDITY
CONC.
mg/l
2700
2280
2420
4600
2600
2560
3050
2700
LOAD
kg. /day
2508
669
710
2362
318
626
1491
726
ALKALINITY
CONC.
mg/l
0
0
0
0
0
0
0
0
LOAD
kg. /day
0
0
0
0
0
0
0
0
TOTAL IRON
CONC
mg/l
740
787
1335
1O33
691
612
750
720
LOAD
Kg. /day
688
231
391
530
84
150
367
194
FERROUS IRON
CONC.
mg/l
728
560
773
672
616
392
507
463
LOAD
kg. /day
676
164
227
345
75
96
?48
125
SULFATES
CONC.
mg/l
3000
21OO
3500
375O
2450
2225
2490
2180
LOAD
kg. /day
2787
616
1027
1926
299
544
1218
586
CALCIUM
CONC.
mg/l
11.4
12.7
17.7
14.7
14.5
14.9
32
100
LOAD
kg. /day
11
3.7
5.2
7.5
1.8
3.6
16
27
MANGANESE
CONC.
mg/l
7.4
7. 1
6.9
6. 1
7.9
9.0
7.8
7.B
LOAD
kg. /day
6.9
2.1
2.0
4.2
1.0
2.2
3.8
2.1
ALUMINUM
CONC
mg/l
8.7
9.2
10.3
9.2
94.5
75.7
88
89
LOAD
kg. /day
8.1
2.7
3.0
4.7
1.2
1.9
4.3
2.4
SPEC.
COND
limhet
2,900
3,150
4,350
4,000
3,050
2.9OO
3.275
3.235
SAMPLING STATION DD-2
SHOFF MINE
DATE
7-30-74
8-20-74
9-O8-74
9-24-74
10-10-74
10-31-74
11-27-74
12-02-74
FLOW
Cu m/min
0.10
O.03
0.09
0.05
0.03
0.02
Drv
Dry
pH
3
UJ
Z
-
2.4
_
2.4
2.2
2.6
_
-
a
4
_l
2.3
2.4
2.7
2.7
2.6
2.5
_
-
ACIDITY
CONC.
mg/t
1700
1408
1460
2500
2000
1820
-
-
LOAD
kg. /day
249
69
178
183
96
44
_
-
ALKALINITY
CONC.
mg/l
0
0
0
0
0
0
-
-
LOAD
kg. /day
0
0
0
0
0
0
_
-
TOTAL IRON
CONC
mg/t
168
116
187
497
271
227
^
_
LOAD
kg. /day
25
5.7
23
37
13
5.5
_
_
FERROUS IRON
CONC.
mg/l
7.84
2.24
s.eo
O
12.3
7.84
-
-
LOAD
kg. /day
1.2
0.1
0.7
O
0.6
0.2
-
-
SULFATES
CONC.
mg/l
1925
1650
2100
2450
1875
2275
_
-
LOAD
kg /day
283
81
257
180
92
56
-
-
CALCIUM
CONC.
mg/l
16.3
14. 1
15.9
19.5
23.9
64.7
_
-
LOAD
kg. /day
2.4
0.7
1.9
1.4
1.2
1.6
-
-
MANGANESE
CONC. ,
mg/l
10.1
9.3
11.2
11.2
13.2
13.2
-
-
' LOAD
kg. /day
1.5
0.5
1.4
0.8
0.6
0.3
-
-
ALUMINUM
CONC
mg/l
4.6
4.2
5.1
3.9
88.5
59.6
_
-
LOAD
kg. /day
. 0.7
0.2
0.6
0.3
4.3
1.5
_
-
SPEC.
COND.
ymhoi
2,700
4,550
4.450
3.900
3,150
2,800
-
-
-------
Table 2 (Cont.)
ABANDONED MINE WATER QUALITY DATA
SHOFF MINE
SAMPLING STATION DD-3
DATE
7-30-74
8-20-74
9-08-74
0-24-74
10-10-74
10-31-74
1 12774
12-O2-74
FLOW
Cu m/min
0.01
O.O3 '
0.02
0.03
0.02
0.02
O O3
0.03
pH
3
ut
£
3.0
_
2.7
3.0
3.6
3 0
3.1
01
2.5
2.6
2,0
2.8
2.7
2.6
2 9
2.8
ACIDITY
CONC,
mg/l
1200
788
940
210O
1300
1120
1BOO
1120
LOAD
kg./day
3,6
39
23
103
32
27
BB
55
ALKALINITY
CONC.
mg/l
0
0
0
0
0
0
Q
0
LOAD
kg. /doy
0
0
0
0
0
0
n
o
TOTAL IRON
CONC
mg/l
25.1
33.8
32.9
28.9
67.7
73.5
69
54.4
LOAD
kg. /day
0.2
1.6
0.8
1.4
1.3
f.9
3.4
2.7
FERROUS IRON
CONC
mg/l
0
5.6
1.12
2,24
2.24
4.48
2.0
1.92
LOAD
kg. /day
0
0.3
0.03
O.1
0.05
O.1
0. 1
0.1
SULFATES
CONC
mg/l
1250
1373
1575
1650
1675
1725
1475
135O
LOAD
kg /day
9.2
67
39
81
41
42
72
66
CALCIUM
CONC
mg/l
16.5
15.5
18.3
17.1
24.7
64.6
152
144
LOAD
kg /day
0.1
0.8
0.4
O.8
0.6
1.6
7.4
7.O
MANGANESE
CONC.
mg/l
10.4
11.5
11.3
10.4
15.8
14.2
1B.4
15
LOAD
kg. /day
0.1
0.6
0.3
0.5
0.4
0.3
n.Q
0.7
ALUMINUM
CONC
mg/l
6.2
5.B
7.1
6.9
95.5
63.2
78
73
LOAD
kg. /day
0.04
0.3
0,8
0.3
2.3
1.5
3. a
3.6
SPEC.
COND
ymhoi
2050
S540
3150
2500
2150
2500
2340
2355
SAMPLING STATION DD-4
SHOFF MINE
DATE
7-30-74
8-20-74
9-08-74
9-24-74
10-10-74
10-31-74
1 1«£7 74
12^32-74
FLOW
Cu m/min
0.01
Drv
Dry
Dry
Drv
Dry
Drv
Drv
pH
3
a
i^
-
-
_
-
_
_
_
-
CO
<
J
2.3
-
_
-
-
_
_
-
ACIDITY
CONC.
mg/l
700
-
_
_
-
_
_
-
LOAD
kg. /day
8.6
-
_
_
.
_
_
-
ALKALINITY
CONC.
mg/l
0
-
_
-
-
_
-
-
LOAD
kg. /day
0
-
_
_
-
-
-
-
TOTAL IRON
CONC
mg/l
12.7
-
_
-
-
_
-
-
LOAD
kg./doy
0.2
-
_
-
-
-
-
-
FERROUS IRON
CONC.
mg/l
10.08
-
_
-
-
-
-
-
LOAD
kg./day
0.1
-
_
_
-
-
-
-
SULFATES
CONC
mg/l
475
-
_
_
_
-
_
-
LOAD
kg /day
5.8
-
_
_
-
_
_
-
CALCIUM
CONC
mg/l
15.2
-
_
_
-
_
_
-
LOAD
kg. /day
0.2
-
_
_
_
_
-
-
MANGANESE
CONC.
mg/l
3.7
-
_
-
_
-
-
-
LOAD
kg./day
0.05
-
_
_
-
_
-
-
ALUMINUM
CONC
mg/l
0.4
-
_
_
_
_
-
-
LOAD
kg./day
0.005
-
_
_
_
_
-
-
SPEC.
COND
i/mhoi
1025
-
_
_
_
_
_
-
SAMPLING STATION DD-5
SHOFF MINE
DATE
7-3O-74
8-20-74
9-08-74
9-24-74
1O-1O-74
10-31-74
1 1-87-74
12-O2-74
FLOW
Cu m/min
0.12
0.05
0.09
0.09
0.02
0.02
0.03
0.002
pH
O
S
_
3.2
_
2.1
3.1
2,8
5.1
2.8
a
<
_i
2.4
2.5
2.8
2.7
2.7
2.6
2.8
2.6
ACIDITY
CONC.
mg/l
900
724
340
1600 '
12OO
1400
1660
1280
LOAD
kg. /day
154
53
103
196
29
34
. 81
3
ALKALINITY
CONC.
mg/l
0
0
0
0
0
0
0
0
LOAD
kg. /day
0
0
O
0
0
0
0
0
TOTAL IRON
CONC
mg/l
170
149
282
410
529
297
366
332
LOAD
kg /day
29
11
34
50
13
7.3
18
O.8
FERROUS IRON
CONC.
mg/l
39.2
90.7
94, 1
80.6
448
134
155
89
LOAD
kg./doy
6.7
6.7
12
9.8
11
3.3
7.6
0.2
SULFATES
CONC
mg/l
1050
11OO
1425
1500
1425
1700
16OO
1410
LOAD
kg /day
1BO
81
174
183
35
42
78
34
CALCIUM
CONC
mg/l
8.9
7.4
10.1
19.1
19.0
3O.5
104
76
LOAD
kg./day
1.5
0.5
1.2
2.3
0.5
0.7
5.1
0.2
MANGANESE
CONC.
mg/l
4.3
5.1
4.5
4.2
6.7
7.9
7.5
6.0
LOAD
kg./day
0,7
0.4
0.5
0.5
0.2
0.2
0.4
0.1
ALUMINUM
CONC
mg/l
3.3
2.7
2.9
0.5
50.9
43.4
50
49
LOAD
kg./day
O.6
0.2
0.4
O. 1
1.2
1. 1
2.4
0. 1
SPEC
COND
ymhos
2225
2975
2950
2475
1800
2300
2620
2490
-------
Table 2 (Cont.)
ABANDONED MINE WATER QUALITY DATA
SAMPLING STATION DD-6 YORKSHIRE NO. 1 MINE
DATE
7-30-74
B-2C
a 24
-74
L-74
10-10-74
1 1-27-74
12-02-74
FLOW
1.00
0.85
0.66
2.96
0.46
4.34
0.56
pH
o
_i
llj
4.9
4.9
5 2
5.B
m
<
_i
3.7
3,8
4.8
4.1
4.7
3.5
5.7
5.6
ACIDITY
CONC.
mg/l
26
16
92
68
32
80
54
116
LOAD
kg. /day
38
2O
83
290
21
499
17
94
101
ALKALINITY
CONC.
mg/l
0
0
10
0
0
0
6
22
0
LOAD
kg. /day
0
0
9.5
O
0
0
1.9
18
TOTAL IRON
CONC
mg/l
52,8
53.8
59.4
51.2
59.0
58.8
55.0
57 2
57.0
LOAD
kg /day
76
66
57
218
39
367
17
46
___ 82
FERROUS IRON
CONC.
mg/l
52.64
40.2
56. 0
63.8
58.2
58.24
53.0
60.0
55.0
LOAD
kg./doy
76
49
54
271
39
363
17
48
SULFATES
CONC.
mg/l
525
400
575
550
500
550
455
440
475
LO»D
kg /day
758
489
548
234O
330
3429
145
355
684
CALCIUM
mg/l
33.4
29.7
34.0
33.9
48.8
52.6
116
12O
117
LOAD
kg. /day
48
36
32
144
32
328
37
97
168
MANGANESE
mg/l
1.5
2.0
, 1.7
1.6
1.9
2.0
2.1
1.8
1.9
LOAD
kg. /day
2.2
2.4
1.6
6.8
1.3
12
0.7
1.5
2.7
CONC
mg/l
7.2
0.5
7. 1
6.9
0
1.3
0.6
0.5
-2^,,,.
LOAD
kg ./day
10
10
6.3
29
0
8.1 ,
0.2
0.4
0.4
SPEC.
COND
pmhoi
850
950
875
9OO
700
SCO
925
975
1.010
U)
SAMPL
DATE
11-2'
-74
r_74
ING STATION DD-7 YORKSHIRE NO. I MINE
FLOW
n n
0.0
Dry
Dry
pH
o
-I
UJ
4.4
m
<
_)
3.2
3.7
ACIDITY
CONC.
mg/l
52
72
LOAD
kg /day
0.4
O.9
ALKALINITY
CONC.
mg/l
0
0
LOAD
kg. /day
0
0
_
TOTAL IRON
CONC
mg/l
1.4
0.3
_
LOAD
kg /day
0.01
0.01
-
CONC.
mg/l
0
_
0
-
LOAO
kg. /day
0
_
0
_
-
[ - _ 1
CONC
mg/l
300
-
525
_
-
_
1 - 1
LOAD
kg /day
2.2
-
_
6.4
-
-
_
1 ^ 1
CONC
mg/l
17.4
-
_
15.3
-
-
-
LOAD
kg /day
0.1
-
_
0.2
-
-
-
mg/l
3.4
-
-
2.9
-
-
-
LOAD
kg /day
0.02
-
-
0.04
-
-
-
-
CONC
mg/l
5.1
-
-
4.9
-
-
-
LOAD
kg. /day
0.04
-
-
0.06
-
-
SPEC
CONO.
limhat
6OO
-
-
725
-
-
.
-------
Shoff Mine
Despite higher total flow rates attributed to its slightly greater area,
pollutant loadings emanating from the Yorkshire Mine were almost
always lower than those from the Shoff Mine. There are several im-
portant factors that contribute to this significant water quality dif-
ference. A large percentage of Shoff Mine headings and workings
were driven to the rise or intercepted by other headings being driven
to the rise. As a result, most of the mine is not inundated, but
drains steadily by gravity to one of several discharge points. Fyritic
coal and fallen roof material is continually exposed to air, ground-
water and moisture condensing from the saturated mine atmosphere.
Three key factors which affect production of acid mine drainage -
accessibility of pyritic material, availability of that material for
reaction, and contact time - all play important roles in explaining
the relative water quality of the two mines. Since all discharges are
located in the southern end of the mine, drainage distance for much
of the mine water is quite long. This extends contact time between
acid-forming materials, and can substantially increase concentra-
tions of mine drainage pollutants. In addition, roof rock above the
workings does not appear to be extremely strong, as evidenced by
collapse of all drifts and surface signs of caving that appear along
much of the coal outcrop. Continual minor roof collapses within
Shoff workings constantly expose fresh, unweathered pyritic materi-
als for reaction. This also substantially increases pollution pro-
duction within the mine.
Yorkshire No. 1 Mine
Mine mapping and associated geologic and water quality information
all suggest that the Yorkshire Mine is largely inundated, with its
overflow point being the drift entry discharge adjacent to Clearfield
Creek. Assuming this statement is true, it is extremely difficult to
explain the source of acid water that periodically discharges from the
slope entry. The mouth of the slope is approximately 21 meters (7O
feet) above the coal seam at that point, and is 11 meters (35 feet)
higher than the drift discharge point. In fact, the slope entry appears
to be at a higher elevation than any portion of the Yorkshire Mine
workings. It is, therefore, highly unlikely that the drainage ema-
24
-------
nating from the slope actually originates in any portion of the mine
workings.
Field observations and the Mine Development map, Figure 4, offer a
feasible explanation for existence of the slope entry discharge. As
the map illustrates, the Lower Kittanning "B" seam has been strip
mined to a point directly adjacent to the slope entry. This seam and
its associated overburden are known acid producers in this area, and
dip of the coal is such that all drainage collected in the strippings
would be channeled to its southern end, where the slope is located.
Available geologic information also indicates that overburden mate-
rial above the Clarion coal, through which the slope was originally
driven, was a relatively unstable roof rock. Since closure of the
Yorkshire Mine, the roof of the slope has probably completely col-
lapsed, blocking the slope and preventing extensive passage of water.
Thus, water channeled through the surface strippings could enter and
pool in the upper portion of the slope, gradually seeping downward
into the mine workings. In wet weather periods, when water pools in
the slope faster than it can seep downward, the discharge appears.
Acidity in this water could result from contact with pyritic materials
in any of several places - Clarion coal overburden in the slope,
Lower Kittanning overburden in adjacent strippings, and refuse from
Yorkshire workings, which is spread throughout the tipple area just
below the slope. This explanation could account for existence of what
appears to be a mine discharge at an elevation far above the predicted
mine pool level.
Assuming the above explanation is valid and that there is only one
actual point of discharge from the Yorkshire Mine workings, hydrol-
ogy and mine drainage in the mine can be accurately assessed. The
Yorkshire Mine's major discharge point is DD-6 (see Figure 2), a
drainage pipe from the mine's collapsed drift entry. As previously
mentioned, water quality in this mine is significantly better than that
observed in the Shoff Mine, although there is a marginal acid pro-
duction problem. The improved water quality is largely due to the
inundated condition of most of the mine, which isolates pollution
forming materials from oxygen required in the acid formation pro-
cess . Neutralization and aeration of some acid may also be occur-
ring fairly close to the point of discharge, as suggested by large
volumes of ferric hydroxide (yellow boy) which intermittently dis-
charge in "slugs."
25
-------
There are several possible explanations for production of the acid
that slightly degrades the Yorkshire Mine water. First, and prob-
ably most important, is the fact that the northernmost tip of the mine
complex is up-dip from the drift entry, and is therefore not inundated.
Thus, in this portion of the mine, pyritic materials are exposed to
air and moisture, and acid mine drainage can form. Since the un-
flooded area is relatively small and contact times are short, rela-
tively little acid forms. Some of this is neutralized, as previously
mentioned, near the mine opening. However, since acid production
occurs fairly close to the discharge point, there may be insufficient
time for neutralization of all acid prior to discharge, A second
factor is the small amount of dissolved oxygen trapped in infiltrating
groundwater which would be available for reaction with acid-pro-
ducing materials. Additional acid waters could be infiltrating down-
ward from surface mines in four of the Allegheny Group coal seams
that overlie Yorkshire workings. Several of these and their asso-
ciated overburden materials are known acid mine drainage producers
in this portion of Pennsylvania.
In summary, water quality data presented in Tables 1 and 2 clearly
shows that there is substantial difference in water quality of tine Shoff
and Yorkshire Mines. The Shoff, with its up-dip development and
unflooded workings, is a major source of acid mine drainage. Dis-
charge quality of the Yorkshire Mine ranges from marginal to
slightly acid, and reflects the hydrologic effectiveness of inundation
as a deterrent to formation of acid mine drainage.
26
-------
Table-3
SUMMARY
ABANDONED MINE SITE CHARACTERISTICS
Yorkshire No. 1 Mine
ShofT Mine
Direction of Operation
Coal Seam
Percent Dip
Period Mined
Area Mined
Percent Extraction
Percent Sulfur In Coal
Percent of Workings Inundated
Maximum Thickness of Overburden
Number of Discharges
Total Flow (cu.m/min.)
Maximum Date
Minimum Date
Average
Total Acidity Load (kg/day)
Maximum Date
Minimum - Date
Average
Total Iron Load (kg/day)
Maximum Date
Minimum - Date
Average
Ferrous Iron Load (kg/day)
Maximum Date
Minimum - Date
Average
Sultate Load (kg/day)
Maximum Date
Minimum - Date
Average
Aluminum Load (kg/day)
Maximum - Date
Minimum - Date
Average
To Dip
Clarion "A" coal
10 South or Southeast
19OO to 1942
22O hectares (54O acres)
80%
3% approximate
90%
90 meters (300 feet)
2
4.34 10/31/74
0.22 11/27/74
1.34
499 10/31/74
17 11/27/74
129
367 10/31/74
17 11/27/74
107
363 10/31/74
17 11/27/74
110
3,429 10/31/74
145 11/27/74
1,009
29 9/24/74
0 10/10/74
7.2
To Rise
Clarion "A" coal
10 South or Southeast
Late 1 SCO's to early 1930's
170 hectares (428 acres)
Variable 35% to 100%
3% approximate
Less than 10%
90 meters (3OO feet)
5
0.89 7/30/74
0.16 1O/1O/74
0.39
2,928 7/30/74
477 10/10/74
1,408
742 7/3O/74
111 10/10/74
365
684 7/3O/74
86 10/1O/74
252
3,265 7/30/74
467 10/10/74
1,398
10.5 11/27/74
3.4 8/20/74
a. 8
27
-------
SECTION V
ACTIVE MINE SITE EVALUATION
MINE LOCATION
Lady Jane Collieries, Incorporated, operates the Stott No. 1 Mine in
Huston Township, Clearfield County, Pennsylvania, about 1.6 kilo-
meters (one mile) south of Penfield (see Figure 5). The mine's
workings lie just east of Bennett Branch Sinnemahoning Creek and
two of its tributaries, Moose Run and Horning Run, beneath approx-
imately 6.5 square kilometers (2.5 square miles) of Moshannon State
Forest. Field offices, drift entry, tipple, treatment facilities and
discharge point for this operation all lie just northeast of Moose Run,
in the southwestern corner of the mine.
From the drift entry, the main heading extends nearly 23OO meters
(7600 feet) to the northeast. The Second North heading then extends
from the main heading to the northwest for 26OO meters (8600 feet),
and actually serves as a main heading for the entire northwest portion
of the mine. Several other drifts also intersect the mine workings,
but are maintained only for ventilation and emergency escape. In
addition, there is a shaft entry to the workings, located north of
Horning Run along the 2nd North heading. This 60 meter (19O foot)
shaft serves as an entryway for mine personnel and supplies.
Specific location, configuration, mine development and general geol-
ogy of the Stott No. 1 Mine are shown in Figure 5.
GEOLOGY
The Stott No. 1 Mine is operated in the Lower Kittanning "B" coal,
which is part of the Pennsylvanian age Allegheny Group. As the
stratigraphic column in Figure 3 shows, rocks overlying the "B"
seam are predominantly sandstones and shales, with two thin, un-
mined coal seams 21 and 51 meters (68 and 166 feet) above. The
"B" coal ranges from 81 to 122 centimeters (32 to 48 inches) thick,
averaging 94 centimeters (37 inches). Analytical values for Lower
28
-------
' .
MOSHANNON
^STATE
- ... ... A
HUSTCXN
GAS WELL
B SHAFT ENTRY
DRIFT PORTAL
C CONVEYOR BELT
W PUMPED WATER
8 COAL OUTCROP LINE
STREAM OR WATER COURSE
MINE AREA (INUNDATED)
iH MINE AREA
1350 STRUCTURE
CONTOUR
V
LADY JANE COLLIERIES INC.
STOTT NO I MINE
0 2000 4000
0.5
FEET
KILOMETER
Figure 5
29
-------
Kittanntng coal produced by the Stott Mine average, on an as received
basis, 3308 Kilogram calories, three percent total sulfur, 12 per-
cent ash and 3.8 percent moisture.
In tints area, the Lower Ktttanntng coal is structurally situated on the
northern flank of the Chestnut Ridge Anticline. Strata here generally
strike about 37 degrees east of North and dtp northwest about 3 de-
grees. No major faults have been identified in the vicinity of the
mine, although minor structural irregularities have been encountered
during mining.
Roof rock in tine mine consists of dark shales and sandy shales. Rel-
atively few fractures are present and good roof conditions exist
throughout most of the mine. The mine floor, or bottom is com-
prised of a relatively soft underclay which, when wet, softens and
presents problems for mobile equipment.
MINE DEVELOPMENT
Stott No. 1 Mine began production in the early 1950's when its drift
entry adjacent to Moose Run was driven into the Lower Kittanning
coal. This drift entry has a surface elevation of 422 meters (1384
feet) above sea level and is located roughly 365 meters (1200 feet)
southeast of the entry to the abandoned Moose Run Mine. From the
drift entry, the Stott Mine was advanced to rise, but not directly
perpendicular to the maximum dip angle, in a northeasterly direction
for about 1293 meters (4240 feet). The coal seam rose approximately
40 meters (130 feet) in that distance, to an elevation of 461 meters
(1514 feet) above sea level. Several south headings were driven
directly up-dip, perpendicular to this main heading, and the south-
eastern portion of the mine was subsequently worked out.
The second phase of mining in the Stott No. 1 Mine extended the main
heading about 730 meters (240O feet) northeast, along the strike of the
coal seam. Panels driven from the north side of this portion of the
main heading were, therefore, down-dip, while panels driven south
were up-dip. While panels in this northeastern portion of the mine
were being worked, 2nd North heading, which serves as the main
heading for the entire northwestern segment of the mine, was driven
down-dip northwest. This heading eventually attained a length of
30
-------
2560 meters (8400 feet) and dropped 125 meters (413 feet) along its
length.
Original mining plans called for 15 left and right headings driven
parallel to strike from the 2nd North heading. Most of the left
headings were eventually driven, and large panels of coal were
worked out along the headings. In addition, it was possible in some
areas to mine as many as three butts, or smaller blocks of rooms
and pillars, from the sides of the panels, thereby increasing their
size. Since headings were driven along strike, differences in coal
elevation only range from 1.2 to 2.4 meters (4 to 8 feet) although
heading lengths average 850 meters (2800 feet). However, differ-
ences in elevation across individual panels, perpendicular to left and
right headings, are as great as 9.1 meters (30 feet).
Although 15 right headings from 2nd North were anticipated during
early mine development, several were never driven. Instead, large
butts were extended from existing mined out panels to permit maxi-
mum extraction of coal where no headings were to be driven. A
recent addition to the mine, located northeast or right of 2nd North
heading, is an auxiliary 2nd North heading, which was driven from
5th Right heading. Coal adjacent to this auxiliary heading will be
mined at some future date.
MINING AND PRODUCTION
Mining Technique
As the Stott No. 1 Mine mapping in Figure 5 shows, current mining
activities are restricted to the 8th Right panel from 2nd North head-
ing. Conventional room-and-pillar mining methods are used, and
coal is transported from the mine by a conveyor belt system. A
working coal face averages about 6 meters (20 feet) in width. This
basic sequence of operations is employed to extract and transport
coal from the active mining area:
1) The roof bolter drills and bolts the roof with bolts of 0.9 to 1.2
meter (3 to 4 foot) lengths. Eight bolts are normally required
for each cut of coal removed.
31
-------
2) After roof control is provided, a face drill is used to drill four
2.4 meter (8 foot) deep holes. The holes are equally spaced
across the 6 meter wide working face.
3) A cutting machine is used to undercut the coal face. The cutting
machine has a 2.7 meter (9 foot) cutting bar which permits un-
dercutting as much as 2.4 meters (8 feet) of coal.
4) After undercutting, the four holes are loaded for multiple shoot-
ing. Two inner holes are fired first with the same time delay.
Shortly afterward, outer holes are fired, using a different delay.
The central blast increases available space in the middle of the
face and provides an opening for coal to be thrown into when outer
holes are fired.
5) A loading machine loads coal into shuttle cars. In the Stott No. 1
Mine, approximately 16 to 18 metric tons (18 to 20 short tons)
are loaded from a single cut.
6) Shuttle cars transport coal to a conveyor or belt feeder. These
cars are 67 meters (22 feet) long and haul 1.8 metric tons (2 short
tons) of raw coal under normal mine conditions. Eight to ten
shuttle car loads are required to transfer all the coal from coal
face to belt feeder. Haulage from face to belt feeder requires
approximately 30 seconds.
7) The belt feeder loads a shuttle car load of coal onto the belt in
about 90 seconds. Coal is then transported out of the mine to the
tipple via conveyor belt system.
Total conveyor haulage distance through the Stott No. 1 Mine is
roughly 3360 meters (11,OOO feet) or over 1.8 kilometers (two
miles). Portions of the belt which convey coal from working face
through 8th Right panel to main haulageway the 2nd North heading -
can be added or removed in 36 meter (120 foot) increments as mining
advances or retreats. The belt is thus always relatively close to the
working face. The main belt then extends 1280 meters (42OO feet)
through 2nd North heading to the mine's main drift heading, and from
there an additional 2070 meters (6800 feet) through the drift entry to
the tipple.
32
-------
Production
Seventy-four people are currently employed (above and below ground)
at the Stott No. 1 Mine. Three shifts are operated - two production
shifts and one maintenance shift. In 1973, production from this mine
totalled 235,000 metric tons (260,OOO short tons), and averaged 190
to 2O9 metric tons (210 to 230 short tons) of raw coal per production
shift. To obtain this production level in a shift, ten to twelve cuts
of coal must be blasted and loaded as previously described.
The 8th Right panel, which is currently being mined, trends parallel
to strike. Therefore, rooms developed on the left side of the panel
are down-dip while those on the right are up-dip. Two production
units are currently in operation, and since both advance and retreat
mining are practiced, each unit periodically operates both to rise
and to dip. For example, as one unit advances to the right from 8th
Right heading, it mines to the rise. Then, as that unit retreat mines,
it is actually mining to the dip.
In order to characterize production achieved by these two units, as
related to their mode of mining (up-dip versus down-dip), production
records were obtained for 1973. This was an average year in terms
of mine operation and production. These records, compiled by the
Mine Superintendent, contain valuable summaries of production from
each of the two units. The following information is presented: tons
of raw coal mined per month, tons of raw coal per man-shift (both
production and maintenance shifts are considered), tons of clean coal
for each production unit per month, monthly activities of each pro-
duction unit, mining conditions encountered, equipment problems,
coal thickness or quality variations, and water conditions. Flow data
representing monthly averages of daily weir flow measurements at
the settling pond effluent point are also included. This 1973 produc-
tion record is presented in Table 4.
Additional production records are also presented for 1967 to charac-
terize the Stott No. 1 Mine's activities prior to passage of the Health
and Safety Act of 1969. As a result of tighter mining restrictions,
coal production generally decreased after passage of the Act. Data
presented for 1967 - 68 production includes most of the same para-
meters already described for the 1973 production log. Production
by the two individual units, however, is not presented separately.
33
-------
Table 4
STOTT NO. I MINE-1973 PRODUCTION RECORD
MONTH
January
February
March
April
May
June
July
August
Seotember
October
November
December
RAN COAL PRODUCTION 1
METRIC TONS
(SHORT TONS)
23,712
(26,143)
19,045
(20,998)
19,947
(21 ,992)
21,012
(23,167)
21,591
(23,805)
23,898
(26,348)
15,806
(17,427)
21,733
(23,961)
20,794
(22,926)
25,324
(27,921)
20,117
(22,180)
20,844
(22,987)
^METRIC TONS^I
\ MANSHIFT A
/SHORT TONS\
\ MANSHIFT/
13.7
(15.2)
11.2
(12.4)
11.2
(12.3)
13.6
(15.0)
13.2
(14.5)
15.9
(17.5)
11.7
(12.9)
13.7
(15.1)
14.7
(16.2)
15.1
(16.6)
13.9
(15.3)
14.7
(16.2)
CLEAN COAL PRODUCTION
METRIC TONS
(SHORT TONS)
UNIT NO. I
9,816
(10,822)
9,439
(10,407)
9,602
(10,586)
9,028
(9,954)
10,965
(12,089)
10,352
(11,414)
7,222
(7,963)
9,651
(10,641)
8,552
(9,429)
11,502
(12,681)
8,955
(9,873)
9,782
(10,785)
UNIT NO. 2
9,O53
(9,981)
8,277
(9,126)
10,220
(11,268)
1O,454
(11,526)
9,739
(10,738)
11,730
(12,933)
7,481
(8,248)
1O, 565
(11,648)
1O, 884
(12,000)
13.12O
(14,465)
10,513
(11,591)
9,519
(10,495)
PRODUCTION
DAYS
22. 0
20.0
21.7
21.0
21.5
20.0
14. 0
19.0
23.0
19. 0
19,0
AVERAGE
MONTHLY
FLOW
n?/ min (cfs)
1.934
(1.138)
1.735
(1.021)
1 .489
(0.876)
1 .671
(0.983)
1.799
(1 .058)
1 .617
(0.951)
1.O16
(0.598)
0.840
(0.494)
O.373
(0.219)
1.610
(0.947)
2.528
(1 .487)
34
-------
Table 4
STOTT NO.I MINE-1973 PRODUCTION RECORD
PROGRESS REPORT
UNIT NO. 1
Advanced to dip
7th L. Panel off 2nd
.North Heading
Advanced to dip
7th l_. Panel off 2nd
North Heading
Advanced to dip
7th U. Panel off 2nd
North Heading
Completed advancing
to dip, retreated BOO
feet to rise . 7th l_.
Panel off 2nd
North Heading
Advancing to dip
7th 1_. Panel off 2nd
North Heading
Advanced to dip
7th L. Panel off 2nd
North Heading
Advanced to dip
7th L. Panel off 2nd
North Heading
Advanced to dip
7th L. Panel off 2nd
North Heading
Advanced to dip
7th L. Panel off 2nd
North Heading
Advanced to dip
7th L. Panel off 2nd
North Heading
Retreat to dip,
then advanced to dip
7th L. Panel off 2nd
North Heading
Advanced to dip
7th U. Panel off 2nd
North Heading
UNIT NO. 2
Advanced to rise
10th R. Panel off 2nd
North Heading
Advanced to rise
10th R. Panel off 2nd
North Heading
Advanced to rise
10th R. Panel off 2nd
North Heading
Advanced to rise
1Oth R. Panel off 2nd
North Heading
Advanced to rise,
then retreated to dip
10th R. Panel off 2nd
North Heading
Advanced to rise
10th R. Panel off 2nd
North Heading
Advanced to rise
10th R. Panel off 2nd
North Heading
Advanced to rise,
then retreated to dip
10th R. Panel off 2nd
North Heading
Advanced to rise
10th R. Panel off 2nd
North Heading
Advanced to rise
10th R. Panel off 2nd
North Heading
Advanced to rise,
then retreated to dip
10th R. Panel off 2nd
North Heading
Retreated to dip
10th R. Panel off 2nd
North Heading
COMMENTS
Unit#1 encountered
some water
New roof bolter
put into service
Flow reduced due
to sludge pumping
No flow due to
sludge pumping
Unit#1 encountered
wet floor
Low flow due to
sludge pumping
35
-------
This 1967 - 68 production data is included in Table 5. An in-depth
evaluation of this data is included in Section VI.
MINE DRAINAGE AND WATER QUALITY
Water is removed from the Stott No. 1 Mine workings by a combi-
nation of pumping and natural gravity flow. As the mapping in
Figure 5 shows, the entire northern portion of the mine has been
developed to the dip. Therefore, lowest portions of the mine, which
have already been worked out, have been allowed to flood to an ele-
vation of 360 meters (11 SO feet) above sea level. The mine pool must
be maintained at this elevation to prevent interference with current
mining activities. Pool level is controlled primarily by two 5O kilo-
watt (75 horsepower) pumps, each with a capacity of 32 liters per
second (5OO gallons per minute). Location of the pumps is also shown
in Figure 5. One is located at the mine pool itself while the other is
located in a sump farther up-dip, where it removes drainage from the
mine before it can enter the mine pool.
Mine drainage is then pumped up 2nd North heading through a 2O cen-
timeter (8 inch) pipe, and then for roughly 760 meters (25OO feet)
along strike in the main heading to a point near 6th South heading.
This drainage then flows by gravity to the main entry. In addition
to the pumped drainage, all other mine workings up-dip from the
main entry, including the large southern section of the mine, also
discharge by gravity.
All water discharged from the mine is treated before it reaches
Moose Run. Discharge flow rates range from 26 to 31 liters per
second (417 to 486 gallons per minute), or about 1.7 cubic meters
per second (one cubic foot per second). During and immediately after
sludge is pumped from the treatment facility's settling basins, flows
are substantially lower, ranging from four to eight liters per second
(64 to 128 gallons per minute). This decrease in flow results from
the increased holding capacity of settling basins immediately after
cleaning.
The mode of formation of the Stott No. 1 Mine's acid drainage is
identical to that discussed for the abandoned Shoff Mine. Large
worked out portions of the mine accumulate infiltrating groundwater
36
-------
and condensation moisture, and drain by gravity toward lower areas.
Gravity drainage permits an extended contact time between the water
and the highly pyritic Lower Kittanning coal, roof and floor mater-
ials. As a result, extremely acid mine drainage is formed and must
be treated to avoid environmental degradation.
The untreated mine drainage pumped from the "B" seam Stott No. 1
Mine is highly acid, and is somewhat similar in quality to water
discharging from the previously discussed Shoff Mine. A typical
water quality analysis is presented below:
pH 2.45
Total Acidity 2800 mg/l
Total Iron 1O09 mg/l
Ferric Iron 742 mg/l
Aluminum 103 mg/l
Sulfate 37OO mg/l
This drainage must be chemically treated in order to comply with
Pennsylvania's effluent standards (pH 6.0 to 9.0, net alkalinity, and
total iron less than 7.0 mg/l). This is achieved using approximately
7.9 kilograms (17.3 pounds) of hydrated lime per 10OO gallons of
drainage. Neutralization is followed by aeration and settling to re-
duce total iron content of the effluent.
In addition to required treatment of mine effluent, there are control
measures employed in the mine to reduce infiltration of groundwater
and subsequent acid formation. One such measure is channelization
of water that does enter the mine to avoid prolonged contact with
pollution forming materials. Channelization is also necessary to
maintain suitable, relatively dry working conditions for mine per-
sonnel and equipment. This can be effectively achieved by employing
small diversion ditches between or along barrier pillars within the
mine.
Infiltration control or reduction is augmented in the Stott No. 1 Mine
by a number of measures. Under certain geologic conditions, drill-
ing for roof bolts may provide access for groundwater to enter the
workings. Where such conditions are encountered in portions of the
mine that will be open or used for extended periods, resin type roof
37
-------
Table 5
STOTT NO. I MINE
1967-68 PRODUCTION RECORD
MONTH
YEAR
September
1967
October
1967
November
1967
December
1967
January
1968
February
1968
March
1968
April
1968
May
1968
June
1968
July
1968
August
1968
RAW COAL PRODUCTION
METRIC TONS
(SHORT TONS)
26,960
(31.929)
29,861
(32.923) '
22,059
(24.321)
27,306
(30, 106)
25,269
(27,860)
17,191
(18,954)
23,46?
(25,871)
24,600
(27, 123)
25,794
(28.439)
24,854
(27,403)
26,223
(28,912)
16,847
(18,574)
METRIC TONS/MAN SHIFT
[SHORT TONS/MAN SHFT)
20.77
(22.9)
20.5
(22.6)
14.3
(15.8)
18.8
(20.7)
17. 1
(18.9)
14.5
(16.0)
16.3
(18.0)
17.1
(18.8)
16.8
(18.5)
16.1
(17.8)
18.7
(20.6)
15.4
(17.0)
PRODUCTION
DAYS
21
22
19
22
20
17.75
21
22
23
20
20
15
38
-------
Table 5
STOTT NO. I MINE
1967-68 PRODUCTION RECORD
PROGRESS REPORT
UNIT NO. 1
Advanced to rise
3rd L. Panel off 4th
South Heading
Advanced to rise
until Oct. 17, then
Retreated to dip
3rd L. Panel off 4th
South Heading
Retreated to dip
3rd I Panel off 4th
South Heading
Retreated to dip
3rd L. Panel off 4th
South Heading
Retreated to dip
3rd L. Panel off 4th
South Heading
Retreated to dip
3rd L. Panel off 4th
South Heading
Advanced to dip
1st L. Panel off 2nd
North Heading
Advanced to dip
1st L. Panel off 2nd
North Heading
Advanced to dip
1st L . Panel off 2nd
North Heading
Retreated to rise
1st L. Panel off 2nd
North Heading
Retreated to rise
1st l_. Panel off 2nd
North Heading until
7/22, then advanced
to dip 2nd L. Panel off
2nd North Heading
Advanced to dip
2nd l_. Panel off 2nd
North Heading
UNIT NO. 2
Advanced to dip
6th L. Panel off 2nd
Nortn Heading
Advanced to dip
6th L. Panel off 2nd
North Heading
Retreated to rise
6th L. Panel off 2nd
North Heading
Retreated to rise
6th L. Panel off 2nd
North Heading
Retreated to rise
6th L. Panel off 2nd
North Heading
Retreated to rise
6th L. Panel off 2nd
North Heading
Advanced to dip
5th L. Panel off 2nd
North Heading
Advanced to dip
5th L. Panel off 2nd
North Heading
Advanced to dip
5th L. Panel off 2nd
North Heading
Advanced to dip
5th L. Panel off 2nd
North Heading
Advanced to dip
5th L. Panel off 2nd
North Heading
Advanced to dip
until 8/2O, then
retreated to rise
5th L. Panel off 2nd
North Heading
COMMENTS
Late Oct. , Unit #2
encountered low
coal, some water
Fire in Unit #1
undercutting saw
Unit #1 utilized
three reduced pro-
duction crews
Unit #2 had limited
working space
Limited working space
poor mining
conditions
Limited working space
for both units in new
locations. Poor roof
Galis 41OO face drill
introduced to Unit # 1
Poor roof required
additional bolting
and timbering
Unit #2 encountered
low coal (less tons per
cut, smaller shuttle
loads, poor maneuvei
ability, increased
b reakdowns)
Unit #2 encountered
low coal. Belts
down for 11.5 hours
(refuse blockage at
transfer points)
3OO man-shifts of
vacation time this
month. Unit #1
encountered wet roof,
bad floor. Unit #2
encountered low coal
39
-------
bolts are utilized. These bolts completely seal the drill hole and
prevent subsequent infiltration, whereas conventional, or expander,
roof bolts do not.
Variation of the mining plan is also utilized as required to avoid
zones of major or minor faulting which could permit increased in-
filtration of groundwater. At the same time, such alterations can
avoid poor roof or floor conditions and lower quality or thinner coal,
thereby increasing productivity.
40
-------
SECTION VI
ANALYSIS OF DOWN-DIP MINING TECHNIQUES
WATER QUALITY
A detailed discussion of the water quality evaluation of abandoned up-
dip versus down-dip underground mines is presented in Section IV.
Water quality data obtained in this study generally confirms prelim-
inary predictions that mining to the dip could be implemented as a
pollution control measure. Sample data for the Yorkshire Mine,
which was developed to the dip, indicated water quality substantially
better than that discharging from the Shoff Mine, which was devel-
oped to the rise. The discharge from the down-dip mine was still,
however, slightly acid in nature and not of acceptable quality. Inun-
dation of the Yorkshire Mine's workings isolated most pyritic materi-
als in the coal, roof and floor, thereby reducing their potential oxi-
dation to form acid mine drainage.
There are several factors which might explain the confirmed forma-
tion of acid in tine Yorkshire Mine. A very small portion of the aban-
doned workings lie above the elevation of the discharge points and are,
therefore, not inundated. Groundwater infiltrating into these up-dip
workings could oxidize pyrite and form acid. Since the point of acid
formation is relatively close to the point of discharge, there may not
be sufficient time to completely neutralize the acid prior to discharge,
even though the mine waters have that potential. A second source of
acid production could be the reaction of dissolved oxygen in the mine
water with the pyritic materials. This could yield limited acid for-
mation and low acidity concentrations in the effluent. Still another
source of acid could be the abandoned surface mines above the York-
shire workings. Spoils of the overlying Lower Kittanning "B" seam,
in particular, are a major source of acid in this area. Acid could be
formed in the strip cuts by surface run-off contact with spoil materi-
als, and could subsequently infiltrate downward into the Yorkshire
workings. Thus, some of the water infiltrating into the mine work-
ings could already be slightly acid in nature.
Despite the minimal pollution evident in the Yorkshire Mine water
quality, it is many times improved over the quality of any of the Shoff
Mine discharges. Since the mines were specifically selected for the
41
-------
many similarities they evidenced, the primary factor controlling wa-
ter quality can only be attributed to the direction of mine development.
Therefore, the water quality benefits to be accrued from utilization of
the down-dip mining techniques can be substantial and the technique
should be given fullest recognition as a pollution control process.
PRODUCTION
Production data obtained for the Stott No. 1 Mine was presented in
Tables 4 and 5 and discussed to some extent in Section V. Inter-
pretation and analysis of that information yields important conclu-
sions concerning effects of down-dip mining on coal production.
Three charts have been developed from available data in an attempt
to visually illustrate production and productivity trends at the Stott
Mine as related to up-dip and down-dip operation. It must be stres-
sed that this data was obtained in only one mine and may not be typ-
ical of all underground mines.
As previously mentioned, data was obtained for a year's production
before and after passage of the Health and Safety Act of 1969. Fig-
ure 6 illustrates a decline in production at the Stott Mine resulting
from implementation of that Act. Also indicated are wide variations
that occur in total monthly production from factors such as: equip-
ment breakdowns, poor roof or floor conditions, excessive water,
and variations in coal thickness or quality.
Trends evident in raw coal productivity data, graphed in Figure 7,
match almost perfectly for both production years with the Raw Coal
Production Chart (Figure 6) previously described. Raw coal produc-
tivity on a tons per man-shift basis also showed a significant decline
after passage of the Health and Safety Act. One reason for decline in
productivity was the required addition of non-production mine person-
nel for health and safety reasons, thereby increasing the number of
personnel while overall production stayed the same or declined.
Figure 7 also attempts to show relationships between advance and
retreat mining and productivity. Valid conclusions can only be
drawn from this chart when both production units are either advancing
or retreating. As the chart shows, this occurred during the first
nine months of 1967-68, but not at all during 1973. Advance mining
generally permitted greater productivity under typical mining con-
42
-------
40,000
35,000
30,000
tfl
I
o 25,000
20,000
15,000
10,000
s
£v?v!
""'""*?£
0 N D
1967
Before
J F
Health
Hgi
i:!*:W
".'.'-'.-'.-'.
;:;::::::7J
:: ;
M A M J J A
1968
and Safety Act
rir:
;
M^pa
r^"
.
-
"-
1~-^_-
~~-^~-
i
~-^~-
_
-
~-^-
~-^-
JFMAMJJASOND
1973
After Health and Safety Act
STOTT NO. I MINE - RAW COAL PRODUCTION
BEFORE AND AFTER HEALTH AND SAFETY ACT
Figure 6
-------
30
.! ,
,1
20
15
M
10
6C
Hr
B ^
,
nr
S
.
__
_iini ^
. __
^ir
__.
0
I9<
N
57
s?s
::^A
D
*
ftiflvi
J
PI
!v,\''v
"M'M'I*
F
I- j
M
i
A
196
_nr
M
.8
^
^
^
&
sScs
^
KW
^
J
1
^
^
^S
^
NTC
^
5w
cS
S^
sss
NN>
SN
j
A
-
J
__
F
M
<^^
y
^
^
>*$
A
-55
M
-
-
J
19-
:
.
j
r3
^
sss
i
5\>
xsS
A
-
-
S
-
"^-
0
^
w
vv
i
\\>
1
N
TO
1
^
i
1
D
Three Productiorv Shifts Were Employed
Rather Then the Usual Two.
STOTT NO.l MINE
RAW COAL PRODUCTIVITY
Figure 7
] Advance Mining
j Retreat Mining
Advance and Retreat
Mining (I unit on each)
-------
ditions. Retreat mining productivity was only comparable for one
month, when skeleton production crews were used on each shift to
permit three production shifts (instead of the normal two).
Greater productivity attributable to advance mining is logical, since
it has several obvious advantages over retreat mining. Retreat
mining generally consists of either splitting or completely removing
pillars left by the advancing unit, there is a smaller volume of coal
available for mining than there was during the advancing phase. In
addition, more equipment maneuvering is generally required during
retreat mining, further decreasing productivity.
The third chart (Figure 8) derived from Stott No. 1 Mine data shows
monthly unit production rates for 1973. This information could not
be graphed for the earlier production year because no unit production
breakdowns were available. Originally the 1973 data was graphed in
an attempt to illustrate trends for mining to the rise versus mining
to the dip. Production Unit No. 1 mined to the dip for most of the
year, while Unit No. 2 worked to the rise; thus, if one mode of oper-
ation yielded increased production, it would be evident as a trend on
this chart. However, no apparent trends can be distinguished, and
it appears that, at least for the Stott No. 1 Mine, mining to the dip
has no substantial production benefits or deficiencies.
Unit No. 2 did produce 5.8 percent more coal in 1973. However,
geologic and mining information reveals that coal mined by Unit No. 2
was thicker by an average of three inches, which would account for a
7.5 percent increase in volume of available coal. To equalize the
monthly production of each unit in terms of coal thickness mined, the
production figures for Unit No. 2 were decreased by 7.5 percent.
These adjusted production figures suggest or estimate production
rates for Unit No. 1 operating to the dip, and Unit No. 2 operating to
the rise, in a coal seam of uniform thickness. Both actual and re-
vised production rates for each unit are shown in Figure 8.
Assuming an equal volume of available coal, production for Unit No.
1 exceeded that of Unit No. 2 by 2.2 percent. The Stott Mine's
mining records, presented in Table 4, also show that Unit No. 2
spent 1.5 more months in retreat mining, with its lower productivity,
than did Unit No. 1. Therefore, when all of these factors are com-
bined and evaluated, production from the two units was approximately
equal for 1973; and mining to the dip was no more or less advanta-
geous than mining to the rise.
45
-------
15,000
j
-.
t -
0
10,000 =
5,000
i
-.
nn
"in
: M
k^^
(Wft
ft 1
^vw
nr:
"^
r~ r
|||
x:>v:;
i^
^^^H
-
Hr:
. .
.
r:?:
1 f\ -
i ii i ij
iiii
'
£$s
-
r^j~
IH]
.
^}^
A r
«»
v |
W£
ITT.
~
n
3£r
iiii
.
"LT"L
2£;
.TIT
HIl
!*'*
~
2^}-
^'-
.
r£r"
IH:
.
i
ITU
PHMW
gg|
WT$'
'^
^-^
ir_r
.
r"~~
«!
*M«J
BBB
~^
in.
.
|:
.
,
u^:
M«
-. -.!!-!
i'/. '::'-
^_
~
,
in:
:tWtf
tV*M
~
t
*
-Ti_r
^m-.
i*i
(1973)
STOTT NO.I MINE
MONTHLY UNIT PRODUCTION RATES
Figure 8
Er£3 Unit No. I (Production To Dip)
llli Unit No.2 (Production To Rise)
mi UnitNo.2 Production Decreased
By 7.5% To Equalize Available
Coal Thickness
-------
There are numerous factors generally unrelated to type of mine de-
velopment which can have adverse effects on production and econom-
ics in underground mines. Combined effects of these factors can
sometimes be so great that they obscure advantages or disadvantages
of mining to the dip. These factors are frequently beyond control of
the operator, and cannot be accurately predicted beforehand. Mine
roof and floor conditions, for example, are extremely important.
Highly fractured roof rock requires increased roof bolting, possibly
augmented by timber cribs, props, crossbars and beams. Additional
temporary supports are required while these permanent supports are
being placed. These roof conditions or fracture zones can also be
associated with excessive groundwater infiltration, resulting in in-
creased pollution formation and pumping requirements. Resulting
wet mine floor conditions, especially where the floor consists of
relatively soft clay, can also adversely affect production. Mobile
equipment and coal shuttle cars may bog down and small pillars and
timbers may sink into the mine floor, causing adjacent portions of
the floor to heave. This infringement on available vertical space in
the workings can hamper production, or even force early closure of
affected portions of the mine.
Geologic and hydrologic conditions can vary sufficiently to hamper
production locally regardless of mining techniques employed. Zones
of foreign mineral veins such as pyrite or clay, decreases in coal
thickness, locally steep dips, and major or minor fault offsets of the
coal seam either decrease the amount of available coal or increase
difficulty of extracting that coal by reducing efficiency of mining,
loading, or shuttle equipment. Some mines are extremely dry, while
others encounter volumes of water sufficient to force closure. The
amount of groundwater encountered in mines is usually dependent
upon a number of factors, all of which relate to infiltration and ca-
pacity of mine workings to intercept that infiltration. Important
factors include orientation and configuration of workings, mine size,
local mining history, mining methods employed, roof and floor rock
conditions, depth and type of overburden, degree of rock faulting and
fracturing, amounts and rates of precipitation, and proximity to
aquifers.
Coal barrier and pillar placement can also have a significant effect
on productivity, since this determines the amount of unmined coal
that must remain. Pillars and barriers vary in size and number
according to mine size, orientation, geologic and hydrologic condi-
47
-------
tions, and depth. Permanent barrier pillars must be left around such
underground obstacles as gas, oil, and water wells, adjacent mine
workings, and heavily used entryways. Large pillars must also be
left beneath surface structures which cannot tolerate subsidence. In
addition, the size of normal pillars increases with depth of over-
burden above the seam.
It can be seen that most production-reducing factors discussed above
do not relate specifically to mining techniques employed. In fact,
most of them are independent variables. Productivity from any
specific mining technique, including down-dip mining, is highly de-
pendent on these factors. Based on this information, and on produc-
tion records reviewed for the Stott No. 1 Mine, no significant pro-
duction decreases have been noted that can be attributed to down-dip
mining.
ECONOMICS
There are numerous factors in the coal mining industry that contrib-
ute to production costs of a ton of coal. Particular attention in this
section is given to economic considerations which may differ for
mines developed to the dip versus mines developed to the rise.
Mention of specific dollar costs incurred at the Stott No. 1 Mine is
avoided to protect competitive interests. Instead, cost items are
presented as a percentage of approximate total production cost per
ton of coal.
Information obtained at the Stott No. 1 Mine is supplemented in this
report with information presented in a recently published United
States Bureau of Mines study of estimated capital investments and
operating costs for bituminous coal underground mines. The small-
est mines evaluated in that study produced about one million tons of
coal annually, substantially more than the Stott No. 1 Mine. How-
ever, much data presented for that production rate is valid and
relevant in describing smaller operations. According to the Bureau's
report, production costs in this tonnage range breakdown in approx-
imately the following percentages:
48
-------
Total labor and supervision
Payroll overhead
Union welfare
Operations supplies
Total power consumption
Indirect and fixed costs
TOTAL 100%
Most primary production cost factors listed above are not affected at
all by utilization of down-dip rather than up-dip mining techniques.
The same mining equipment is utilized; thus, there is no change in
number of mine personnel. This means total labor, supervision,
payroll overhead and union welfare remain unchanged. Since actual
coal extraction techniques do not vary substantially, operations
supplies and indirect and fixed costs also remain approximately the
same. Thus, the parameter showing greatest degree of variation
reflecting production economics of up-dip versus down-dip mining
is power consumption.
The portion of production costs relating to power consumption rep-
resents the most critical cost variable in a comparative evaluation of
down-dip and updip mining. As underground mines increase in size,
power consumption percentage of total production costs declines
below the 3 percent shown above. Total power costs in the Bureau
of Mines study ranged from 11 to 17 cents per metric ton (12 to 19
cents per short ton). One active mine using continuous miners and
rail haulage reported power consumption costs of $0.56 per metric
ton ($0.50 per short ton), but even that represented only 4 or 5 per-
cent of the total production cost.
Since power consumption is a key variable in assessment of down-dip
mining economics, a closer examination of the relative power con-
sumption percentages of various pieces of mining equipment is in
order. Equipment utilizing substantial amounts of power includes
continuous miners, undercutting saws, loading machines, shuttle
cars, roof bolters, belt feeders, secondary conveyor belt systems
used in headings, gathering pumps, ventilation fans, rock dusters,
and shaft and slope hoists. There are other smaller consumers of
electrical power too numerous to mention. The Stott Mine, which is
smaller than any of the mines evaluated in the Bureau of Mines'
49
-------
.eport, had a slightly higher power consumption cost percentage.
lectrical costs at the Stott Mine account for about five percent of
5tal production costs. The following table of equipment and power
onsumption percentages provides a relative comparison of electrical
osts for normal operations at the Stott No. 1 Mine.
Tables
STOTT NO. 1 MINE
POWER CONSUMPTION BREAKDOWN
Percentage of Total Mine
Equipment Hours/Week Power Consumption
^/lain belt drives 8O 37.9%
Ventilation fan 168 16.6
v\ain pumps 75 11.1
Drift to stacker belt drive 80 8.6
Side belt drives 8O 6.3
Treatment plant 168 5.0
Roof bolter 30 3.6
Tipple 35 3.5
Loaders 3O 1.8
Shuttle cars 40 1.4
Face drill 4O 1.2
Undercutting saw 10 1.0
Miscellaneous - 2.0
100.O%
50
-------
As Table 6 shows, haulage of coal constitutes by far the single
largest power cost, despite the fact that the Stott Mine only produces
coal on 2 shifts per day. Haulage during those two shifts accounts
for 53 percent of all power consumed by the mine, or about 2.5 per-
cent of total production cost. The Stott Mine accomplishes haulage
exclusively through primary and secondary conveyor belt systems.
This means of coal transport can, depending on mine conditions, be
much more efficient than rail haulage, and has gained almost ex-
clusive use in newer underground mines. Belts have a significant
advantage over rails in that rails are restricted to grades of less
than three degrees, while belts can be efficiently utilized where
grades are as high as 14 to 16 degrees.
Terminology used in discussions of conveyor haulage in up-dip and
down-dip mines is potentially confusing, thus a brief explanation is
presented. In a down-dip mine, where development progresses
toward lower elevations and the coal face is down-grade from the
mine entry, conveyors must haul coal up-grade. The opposite is
true of mines developed to the rise, where coal must be hauled
down-grade from the working face to the mine entry.
Unfortunately, cost and power consumption figures for the Stott
Mine's two production units could not be separated to the extent
necessary to effectively evaluate coal haulage up-dip versus down-
dip. To compensate for this data gap, a conveyor belt system manu-
facturer was contacted to determine how belt system operating costs
were computed during planning and development of mine haulage
systems. Information obtained from this source was invaluable for
computation of conveyor haulage costs under various conditions.
There are a number of variables which must be considered when
attempting to determine belt haulage costs. Factors such as antici-
pated production and haulage rates, layout of the mine workings,
angle of grade to the rise or dip, belt segment length, and the width,
composition, and velocity of the belt are all important in computing
horsepower, energy consumption and costs of belt haulage. Table 7
includes many of these variables, and defines their effects on up-
grade and down-grade conveyor haulage costs per metric ton.
The primary purpose of Table 7 is evaluation of unit cost differentials
for up-grade conveyor haulage, as would be required in a mine devel-
oped to the dip, and down-grade haulage, which would be employed in
an up-dip mine. To permit this, a constant belt length of 305 meters
51
-------
Table 7
CONVEYOR BELT OPERATION COSTS
UP-GRADE VS. DOWN-GRADE HAULAGE
BELT
WIDTH
61 cm,
(24 In.)
76 cm.
(30 In.)
HAULAGE
RATE
KKG/HOUR
(SHORT
TPH)
45.4
(50)
80.7
(100)
ise'.i
(150)
181.4
(200)
226.8
(250)
272.1
(300)
317.5
(350)
362.8
(400)
408.2
(450)
453.5
(500)
HAULAGE
DIRECTION
Rise
Dip
Rise
Dip
Rise
Dip
Rise
Dip
Rise
Dtp
Rise
Dip
Rise
Dip
Rise
Dtp
Rise
Dip
Rise
Dip
ELEVATION CHANGE IN 1000 METERS
(DEGREES DIP OF BELT)
5m
(0.30°)
0.46
0.44
0.27
0.25
0.21
0.19
O.17
0.16
0.15
0.14
0.16
0.14
0.15
0.14
O.14
0.12
0.13
0.11
0.12
0.11
10m
(0.57°)
0.46
0.44
0.28
0.25
0.2:1
0.18
0,18
0.15
O.16
0.13
0.16
0.13
0.15
0.12
0.15
0.11
0.14
0.11
0.13
0.10
15m
(0.86°)
0.48
0.42
0.29
0.24
0.22
0.17
0.19
0.14
0.17
0.12
0.18
0.13
0.16
0.11
0.16
0.10
0.15
0.1O
O.14
0.09
20m
(1.15°)
0.48
0.42
0.30
0.23
0.23
O..17
0.2O
O.14
0.18
0.12
0.19
0.12
0.17
0.11
0.17
0.10
0.16
O.O9
0.15
0.09
25m
(1 .43°)
0.50
0.42
0.31
0.22
0.24
0.15
0.21
0.13
0.19
0.11
0,19
0.11
0.18
0.10
0.17
0.09
0.17
0.08
0.16
0.08
30m
(1.70°J
0.50
0.42
0.32
0.22
0.25
0.15
0.22
0:12
0.20
0.10
0.20
0.11
O.19
0.09
0.18
0.09
0.18
0.08 .
0.17
0.07
40m
(2.30°)
0.52
O.40
0.33
0.20
0.27
0.14
0.24
0.11
0.22
0.08
0.22
0.09
0.21
O.O8
0.20
0.07
O.2O
0.06
0.19
0.06
5Om
(2.86°)
0.54
0.38
0.35
0.19
0.29
0.12
0.26
0.09
0.24
0.07
0.24
0.08
O.23
0.07
O.22
0.05
0.22
0.05
0.21
0.04
6Om
(3.43°)
0.56
0.36
0.37
0.17
0.31
0.11
0.28
O.08
0.26
O.O6
O.26
0.06
0.25
0.05
O.24
0.04
0.23
0.04
0.23
0.03
80m
(4.57°)
0.60
0.34
0.41
0.15
0.35
0.08
0.31
O.05
0.29
0.03
O.30
0.07
O.29
O.02
0.28
0.02
0.27
0.01
0.26
0.01
100m
(5.70°)
0.64
0.30
0.45
0.12
O.38
0.05
0.35
0.02
0.33
O.01
0.33
O.O1
*
*
*
*
125m
(7.12°)
0.68
0.28
0.49
0.08
0.43
0.02
*
*
*
*
*
*
*
150m
(8.50°)
0.74
0.24
0.54
0.05
*
*
*
*
*
*
*
*
NJ
* Down-dtp haulage yielded negative values, which are not valid. In such cases, the length of the belt
segment would be reduced to permit computation of accurate costs.
NOTE: Costs are In cents per kkg of coal hauled per 305 meter belt segment.
-------
(1000 feet) and velocity of 122 meters per minute (400 feet per
minute) were assumed. Ten different production rates, ranging
from 45.4 to 453.5 metric tons per hour (50 to 500 short tons per
hour) were considered, as were two different belt widths - 61 and
76 centimeters (24 and SO inches). A range of slope or grade angles
between O.3° and 8.5° is evaluated, and costs per metric ton for
belt transporting coal up and down those grades are computed.
Table 7 exhibits several noteworthy trends concerning costs of
utilizing conveyor belts to transport coal. As would be expected,
costs for haulage up-grade increased with greater slope angles and
the opposite was true for down-grade haulage. Second, the cost per
metric ton for transporting coal over a constant belt length at a con-
stant velocity was extremely low (always less than one half cent per
ton) regardless of other mine conditions. The table also shows that
use of larger belts and higher haulage rates can significantly lower
the cost per metric ton of belt haulage.
This table also illustrates cost variations between up-grade and
down-grade haulage over a predetermined slope angle. The pei
centage variation was computed for each pair of up-grade and down-
grade unit haulage costs evaluated. These percentages are shown
in Table 8 and are graphically illustrated in Figures 9 and 1O. The
table and figures show percentage increase in haulage costs attrib-
utable to upgrade rather than downgrade coal haulage ranges from
a few percent in shallowly sloping settings to several hundred percent
in steeper slopes. However, the table and the figure can be some-
what misleading if not carefully considered. While cost increases of
30O percent or 400 percent for up-grade haulage appear quite sub-
stantial, the actual variations in haulage costs are generally much
less than one half cent per metric ton.
To summarize the conveyor haulage statistics that have been pre-
sented here, it appears the economic impact of up-grade versus
down-grade coal haulage is relatively minimal. There may be a
significant haulage cost increase in terms of percentage of down-
grade haulage, but unit costs for both modes of haulage are so low
'generally from O.1 to O.5 cents per metric ton) that these increases
ire insignificant. Thus, direction of belt coal haulage in any mining
situation, including mining to the dip, does not appear to be a sig-
nificant economic factor.
53
-------
Table 8
PERCENTAGE COST INCREASES IN UP-GRADE
OVER
CONVEYOR HAULAGE DOWN-GRADE
BELT
WIDTH
61cm,
(24 In.)
76 cm,
(30 In.)
HAULAGE
RATE
KKG/
HOUR
(SHORT
TPH)
45.4
(50)
90.7
OOO)
136.1
(150)
181.4
(200)
226.8
(250)
272.1
(300)
317.5
(350)
362.8
(400)
408.2
(450)
453.5
(500)
ELEVATION CHANGE IN 10OO METERS
(DEGREES DIP OF BELT)
5m
(0.30°)
4.5
5.9
10.5
6.3
7.1
14.3
7.1
16.7
18.2
9.1
10m
(0.57°)
4.5
12.O
16.7
20.0
23.1
23.1
25. 0
15.4
27.3
30.0
15m
(0.86°)
14.3
20.8
29.7
35.7
41.7
38.5
45.4
60.0
50.0
55.6
20m
(1.15°)
14.3
3O.4
35.3
42.9
50.0
58.3
54.5
70.0
77.8
66.7
25m
(1.43°)
19.0
40.9
50.0
61.5
72.7
72.7
80.0
88.9
112.5
100.0
30m
(1.70°)
19.0
45.5
66.7
83.3
100.0
81.8
111.0
100.0
125.0
142.9
40m
(2.30°)
30.0
65.0
92.9
128.6
175.0
144.4
163.0
186.O
233.0
216.7
50m
(2.86°)
42.1
84.2
141.7
188.9
242.8
200.0
229.0
34O.O
340.0
425.0
60m
(3.43°)
55.6
117.6
181 .8
250.0
333.0
333.0
400,0
500.0
475.0
667.0
80m
(4.57°;
76.5
173.3
337.5
520.0
867.0
329.0
1350
13OO
2600
250O
100m
(5.70°)
113.3
275.0
66O.O
1650
3200
3200
*
if
*
*
125m
(7.12°)
142.8
512.5
2000
*
*
*
*
*
*
*
150m
(8.50°)
208.3
980.0
*
*
*
*
*
*
*
*
Cn
.to.
* Down-
-------
n
40
80
120
160
200 240
280
320
360 40O 440
480
520
560 60O
Percent
640
PERCENT INCREASES IN COST PER KKG
OF 63 cm.BELT COAL HAULAGE UP-GRADE
VERSUS DOWN - GRADE
Figure 9
-------
10
8
« ir
00 o
jr F
° « 5
« Q,
o, »
< S1
4
300 tph
80
120
160
200 240 280 320 360 400 440 480
PERCENT INCREASES IN COST PER KKG
OF 76 cm. BELT COAL HAULAGE UP-GRADE
VERSUS DOWN-GRADE
520
56O 600
Percent
640
Figure 10
-------
Pumping is a second key parameter in underground mine power
consumption that could be seriously affected by employment of down-
dip mining techniques. All of the workings in any mine drain by
gravity to the lowest point. In most down-dip mining plans, that
lowest point is located in the active portion of the mine. Since there
is no discharge point there, and all mine water is draining toward the
coal face, continuous or at least intermittent pumping is frequently
required to maintain a dry working area.
Pumping is a highly variable parameter, dependent on geologic and
hydrologic conditions, type of equipment utilized, and the mining
plan. A mine developed exclusively to the rise would obviously have
a distinct advantage over one developed to the dip, since the up-dip
mine would drain largely by gravity to the discharge point. Pumping
would still be required in some actively worked panels, but equipment
and time expenditures would be relatively small. In an advancing
down-dip mine, where virtually all water must be removed to permit
continued production, pumping requirements are totally dependent
upon mine size, geology and hydrology.
Another factor which may have a minimal effect on a mine's pumping
requirements is utilization of continuous miners. These machines
utilize a low volume water spray to control dust at the coal face. If
the mine is operated in an acid-producing coal seam, all drainage
within the workings may be too highly acid and corrosive to use in the
sprays. Additional water must then be pumped from the surface to
the continuous miners for spraying purposes, thereby increasing the
volume of water that must be eventually pumped. Depending on the
number of continuous miners in operation, this additional volume of
water may be very small or fairly large.
Information available for pumping is general in nature, however, and
can lead to no specific conclusions. It can only be stated that pump-
ing costs may be substantially increased by down-dip development,
but will probably not reach the point where they adversely affect pro-
duction economics. The Stott Mine might exemplify this, since it
can be considered an average mine in terms of amount of pumping
required. Even here, energy consumption costs for pumping rep-
resent less than one percent of the mine's total production cost.
Many of the other power-consuming processes or pieces of equipment
listed in Table 6 will not be affected by utilization of down-dip rather
57
-------
than up-dip mining procedures. Undercutting saws, roof bolters,
face drills, loaders and shuttle cars all operate equally well in any
grade or direction of grade normally encountered during mining. No
changes in mining operations, their sequence or numbers of equip-
ment operating personnel are required. Ventilation fans and tipple
facilities will also maintain the same power consumption levels
regardless of the manner of mine development.
One of the most critical mining cost factors and potential advantages
of down-dip mining is realized only after mine closure. As pre-
viously mentioned, closure of mines developed to the rise requires
expensive sealing of entire mines or many segments of some mines.
The ability to effectively seal a mine and prevent discharges is deter-
mined by three factors: 1) hydraulic head that will ultimately affect
the seal; 2) integrity of the coal adjacent to the seal; and 3) strength
and integrity of coal barriers surrounding the mine. Mine seals
frequently cannot be constructed to dependably withstand more than
6.1 to 7.6 meters (20 to 25 feet) of hydraulic head. Many undei
ground mines in Appalachia, however, have total elevation changes
many times greater than this. Such mines would have to be sealed
in segments, each segment having less than the maximum permis-
sible head. This can be extremely dangerous because failure of any
one of the multiple seals could place an undue stress on consecutively
lower seals, resulting in their failure and subsequent failure of the
entire sealing operation.
One of the principal problems in seal construction lies not with the
seal, which can be designed to withstand almost any amount of pres-
sure, but with integrity of coal adjacent to the seals. As a result of
mining, this coal is frequently in a fractured condition, and seals
cannot easily be tied to or interlocked with the mine walls. The
result is slowly increasing seepage around the seal as hydraulic
pressure increases.
The coal barrier itself must also be of sufficient strength to maintain
anticipated pressures. Where a mine is developed to the rise, the
barrier on the down-dip or entry side of the mine will be a portion of
the coal outcrop. If mine workings came too close to the outcrop, as
was frequently the case when "grass roots" mining was common
years ago, the outcrop coal barrier can be substantially weakened,
rendering it incapable of withstanding any great amount of pressure.
This, too, can result in failure of a mine sealing project, either by
58
-------
steadily increasing seepage rates or by complete outcrop failure.
Coal barrier integrity is no longer a major sealing factor in some
of the states which have some form of barrier pillar requirements.
However, it is not uncommon for such requirements to be excessive.
Excessive barrier restrictions serve no valid safety purposes, and
only tie up valuable coal reserves. Thus, poorly established re-
quirements can result in an economic and energy loss.
When mines are developed to the dip, the first of these problems
concerning strength of tine seal and its bond to mine walls, are rela-
tively unimportant. Since mine entries are on the up-dip side of the
workings, they will be subjected to very little or no hydraulic head
following closure. The question of coal barrier integrity becomes
extremely important, however, because the workings will naturally
inundate. If the low side of the mine workings is adjacent to a coal
outcrop or the workings of another active or abandoned mine, care
must be taken to leave enough coal in the barrier pillar to withstand
anticipated heads upon closure. Since all coal left in the barriers is
potential profit which can never be realized, it is also important to
assure that only the barrier thickness required to safely maintain
mine inundation remains unmined. This allows a maximum amount
of coal to be extracted during down-dip mining without reducing safety
or effectiveness of mine inundation.
In the past, both industry and government were concerned more with
barrier pillar widths for reasons of safety, mine subsidence and
property rights. Until recently, little or no consideration has been
given to the importance of barrier pillars in maintaining post-mining
inundation, which in turn minimizes water pollutant production and
preserves environmental quality. Attention has only recently been
focused on evaluation of barrier pillar requirements.
The Commonwealth of Pennsylvania has been a leader in assessing
importance of adequate barrier pillar thicknesses. Pennsylvania's
current barrier pillar restrictions, which are actually "rules of
thumb" rather than legislative doctrines, are based primarily on
predictions from past mining experiences. Prior to 1967, this
State's barrier pillar widths were determined by thickness of the
seam mined, but this proved inadequate when hydraulic heads ex-
ceeded 15 meters (5O feet). Post-1967 regulations require mines
developed to the dip maintain at least a 15 meter (50 foot) barrier
59
-------
along the coal outcrop and between adjacent workings. Mines devel-
oped to the rise are governed by regulations which vary according to
the anticipated hydraulic head. Hydraulic heads of less than 30
meters (1OO feet) require at least one half meter (1.5 feet) of barrier
pillar per foot of head. Where anticipated hydraulic heads exceed
91 meters (30O feet), mining is generally discouraged.
Several other states also have barrier pillar restrictions, some of
which are based upon Pennsylvania's "rule of thumb." Requirements
in West Virginia and Maryland are similar to those in Pennsylvania.
Kentucky permits mining to within 7.6 meters (25 feet) of adjacent
property lines or mines. Alabama requires a 152 meter (5OO foot)
barrier pillar around each mine, regardless of conditions or size.
Virginia requires a 61 meter (20O foot) property line barrier only
where the adjacent property is mined. Tennessee and Ohio have no
definite barrier pillar requirements.
The wide range of variability in these regulations points to the fact
that the problem of assessing barrier pillar conditions and deter-
mining thickness requirements has not yet been solved. With the
increase of down-dip mine development for pollution control, this
problem will require extensive consideration, hopefully culminating
in development of an effective formula or equation for predicting ade-
quate but not exaggerated barrier pillar requirements. Dependable
regulations would maximize economics of mining to the dip while
assuring that environmental degradation is prevented.
HEALTH AND SAFETY CONSIDERATIONS
Down-dip mining is not a new and untried mining technique, but one
which has been locally implemented for many years. Therefore, the
health and safety aspects of this technique have been well established,
eliminating need for speculation. There is no significant difference
in any aspect of down-dip mining as opposed to up-dip mining. The
same equipment, personnel, coal extraction methods and haulage
techniques are employed. The only alteration of normal mining pro-
cedures may be pumping increases required to minimize water im-
poundment at the coal face, but this is highly variable and does not
affect health and safety. In addition to these points, both up-dip and
down-dip mining must remain in compliance with the stringent,
60
-------
effective requirements of the Health and Safety Act of 1969. Thus,
there are no noteworthy health and safety aspects of down-dip mining
that could adversely affect its implementation.
NATIONAL IMPACTS
National impacts of mining to the dip rather than to the rise are dif-
ficult to assess in terms of water quality improvements, because
mine drainage pollution is not really a nationwide problem. Pro-
duction of mine drainage pollutants (acid, iron, sulfates, etc.) is a
regional and local phenomena, dependent upon regional character,
local geology, specific coal seams mined and extent of mining. The
most significant pollutant attributable to coal mining is acid mine
drainage, which is widespread only in the Eastern or Appalachian
Coal Province. States in this region which are faced with degrada-
tion of streams by acid mine drainage include Pennsylvania, Ohio,
Maryland, Virginia, West Virginia, Kentucky, Tennessee, and
Alabama. There are also local acid pollution problems in other por-
tions of the Nation, as well as non-acid areas in which iron or other
mine drainage pollutant concentrations are above acceptable levels.
In all of these cases, only a few of the coal seams present are pol-
lution sources.
Regional structure of the Nation's various coal basins is also a rele-
vant factor in determining national impacts. Much of the coal west
of the Appalachian Basin dips very shallowly or is flat-lying. In
such areas, mines may extend in any or all directions with no con-
sideration for the minimal dip that does exist. Down-dip mining is,
therefore, only a valid mining technique in those coal basins in
which dips of the seams exceed one or two degrees.
The extent and type of mining conducted in an area is second in im-
portance only to the chemical quality of coal and overburden materi-
als in determing acid production extent and severity. Highly pyritic
coal seams which have not been exploited by mining are seldom
significant sources of mine drainage. Numerous studies conducted
in the Appalachian coal field have attributed 7O to 90 percent of all
acid mine drainage pollution to abandoned underground mines. The
vast majority of these mines were originally developed to the rise
to permit gravity drainage and coal haulage.
61
-------
Pollution significance of previously abandoned underground mines
cannot, however, be construed to imply that underground mines
abandoned now or in the future will be similar pollution sources.
Recently proposed effluent limitations guidelines for the coal industry
require that mine operators be held permanently responsible for the
quality of effluent from their mines, during mining and after aban-
donment. They must meet certain effluent standards regardless of
the mining techniques they employ. Thus, in terms of water quality
improvement on a national scale, employment of down-dip mining
will have little effect, because all mine discharges must be of ac-
ceptable quality.
Down-dip mining will, however, be of major economic importance to
mine operators who, as a result of these proposed effluent limita-
tions, are faced with the prospect of perpetual treatment of dis-
charges from their abandoned operations. Such permanent treatment
will be absorbed as a production cost increase in the operator's
other active operations, and will therefore have a significant national
impact. The U. S. Environmental Protection Agency's recently
prepared Development Document for Effluent Limitations Guidelines
and Standards of Performance for the Coal Mining Point Source
Category presents treatment cost data for several of the coal indus-
try's best acid mine drainage treatment plants all of which employ
lime neutralization. That study indicates costs of mine drainage
treatment are dependent on several factors* Construction costs are
largely determined by the volume of water that will be treated. Fa-
cilities that must be included in computation of construction costs
for a treatment plant are land, holding basin, control building, lime
storage, lime and feed mixer, aeration facilities, settling basins,
fencing and roads, sludge disposal equipment and basins, instruments
and electrical apparatus, and pumps.
Since most treatment plants employ earthen settling basins in their
treatment process, land requirements can become very significant.
Treatment facilities are generally confined to land requirements of
less than ten acres. Even this may present problems where under-
ground mines are concerned, since surface ownership is frequently
confined to the mine access areas. However, land acquisition is
generally not a significant aspect of treatment facility cost.
Disposal of sludges produced in treatment of mine drainage is an
increasing problem which must be considered. Recently constructed
62
-------
treatment plants are providing settling basins with capacities of many
millions of liters and provide sludge storage for several years. In
some plants, sludge is intermittently removed for disposal in aban-
doned portions of the underground mine workings.
In addition to plant construction costs, treatment costs must also be
considered. Costs in the previously cited EPA document ranged from
$O.O3 to $0.11 per thousand liters ($.11 to $.40 per thousand gallons)
regardless of the acidity concentration. Lime is the most commonly
used neutralizing agent, but other chemicals that can be used include
limestone, soda ash and caustic soda. Limestone, the raw material,
is readily available for production of lime; however, there is pres-
ently a tight market for availability of lime due to closing of several
plants for air pollution problems. Availability of lime and other al-
kalies is a factor that could significantly increase costs of future
treatment of underground mine discharges.
Even where all acid has been removed from the mine effluent by
treatment, water quality may be unacceptable for certain applica-
tions. Such parameters as sulfates and total dissolved solids cannot
be economically removed by any treatment process, while others
such as hardness and total suspended solids may actually be in-
creased during treatment of the mine drainage.
This extensive discussion of the complexities of acid mine drainage
treatment is to amplify potential advantages of the down-dip mining
technique. Water quality evaluation in this study has revealed that
mining to the dip can yield significantly improved effluent water
quality after mine closure. Natural inundation decreases pyrite
oxidation and less acid is formed. Even if the effluent from down-
dip mines is still slightly acid, as was the case with the Yorkshire
Mine, the subsequent costs of perpetual treatment of that discharge
are substantially lower than they would be if the same mine had been
developed to the rise.
Thus, implementation of down-dip mining in major acid-producing
regions could have significant regional and national impact on pro-
duction costs. The only noteworthy production cost increases attrib-
utable to down-dip mining are pumping costs which, as previously
mentioned, are so dependent on local geology and hydrology they
cannot be predicted on a regional scale. Post-mining treatment cost
savings attributable to down-dip mining are impressive, when com-
63
-------
puted on a perpetual basis, and will be more than sufficient to offset
any increased pumping costs. Then, National impact of increased
utilization of the down-dip mining technique would actually be
favorable rather than adverse.
FEASIBILITY AND APPLICABILITY
Feasibility and applicability of the down-dip mining technique have
already been established in numerous applications under various
mining conditions. These applications have revealed that there are
no real technological limitations to use of the technique, although
production economics must be evaluated at each potential minesite.
Evaluation of the technique in this study has also revealed there are
no significant production cost increases attributable to actual coal
extraction or haulage processes. The study shows that pumping is
the only significant mining variable that can be adversely affected by
down-dip mining. Increased pumping can lead to increased water
treatment, and can severely effect production economics at a mine-
site. However, hydrological and geological conditions that relate
to groundwater infiltration vary on a local basis and must be eval-
uated at each proposed minesite, regardless of mining technique
being employed. To counter any increased water handling attrib-
utable to down-dip mining, several alternative mining procedures are
available.
Since down-dip mining is not an entirely new or experimental tech-
nique, its range of applicability has already been fairly well estab-
lished through actual use. Normal down-dip mining procedure has
been to enter a coal seam from an up-dip coal outcrop and expand
from there toward the dip. As the mine develops to the dip, all
infiltrating water entering mined out areas, which are up-dip from
the advancing coal face, drains by gravity toward the working face.
Depending on local conditions, extensive pumping may be required
at the coal face to permit mining to continue. As mining proceeds,
roofs of older abandoned portions of the workings may settle and
fracture, causing increased infiltration and necessitating even more
pumping.
There are several alternatives or options to this conventional tech-
nique that can help reduce pumping requirements. The first of these
64
-------
is merely a modification that is already employed in many mines.
Main headings and general trend of workings is still developed toward
the dip, with all secondary headings developed parallel to the strike.
Rooms or panels driven from these secondary headings are then
oriented either to the rise or to the dip. If down-dip panels are
developed first, water to be handled will be only that in the immediate
vicinity of the active face. Infiltrating groundwater can be allowed to
accumulate in these completed panels as up-dip panels are being
developed. Local inundation of these down-dip panels can eliminate
some flow of mine waters to the lowest portions of the workings, and
thus reduce pumping requirements.
Another alternative down-dip mining technique also has potential to
substantially reduce production costs by decreasing pumping and
treatment requirements. This involves development of a down-dip
mine in such a manner that actual mining would generally advance to
the rise. To accomplish this, the main heading or headings are
initially driven from the up-dip coal outcrop to the lowest anticipated
point in the proposed mine. Mining would then expand from this
lowest point, at the end of the main heading, and advance to the rise
pulling pillars as mining progressed. Initial full-scale production
would then occur from the down-dip workings and progress to the
rise.
Use of this alternative has several potential advantages and disadvan-
tages which would have to be weighed with local minesite conditions
to determine coal production economics. Drainage reaching the
newly developing lower portions of the mine would be only that which
infiltrated through roof and walls of the main heading, and would,
therefore, be minimal in volume. This would also be the case in a
normally developed, new down-dip mine, where the relatively small
active and mined out areas permit only limited infiltration. Thus,
little or no advantage in pumping is realized in earlier stages of
mining.
Pumping advantages of this alternative down-dip mining technique are
increasingly realized as the mine expands, pillars are split or pulled,
and portions of the workings are abandoned. In many down-dip op-
erations, these abandoned areas may be higher in elevation than the
active working areas, and therefore, infiltrating water drains by
gravity to the active face, where increased pumping is required. If
mining is advancing to the rise from the mine's lowest point, infil-
65
-------
trating water can be allowed to inundate low portions of the mine
after pillars are removed or split. Thus, as mining is progressing
to the rise, the extent of inundation of lower worked out areas is
also increasing, and little or no pumping of water from the workings
is required. Adoption of this mining alternative would decrease
pumping costs, thereby decreasing overall production costs.
Use of belt haulage in these mines will also require some modifica-
tion, but this should not affect overall economics of mine operation.
In normal down-dip mining, main haulage belt segments are added
as mining progresses. This proposed up-dip development technique,
however, would require the entire length of main haulage belt be
constructed to the lower limits of the mine prior to initiation of full
scale production in that area. Therefore, initial investment in mines
with this type of development plan must include all required main belt
segments, since they must all be in position before full scale pro-
duction can begin. Then, as mining proceeds to the rise, main haul-
age distances decrease and belt segments must be removed. The
number of belt segments required would be the same, but the se-
quence of addition or removal of those segments would be altered.
This would cause an increase in haulage and production costs from
the newly developing operation, but both costs could be expected to
decline as mining progressed. Thus, the average haulage production
costs over the life of the mine would not be significantly altered by
implementation of this alternate down-dip mining technique.
A possible disadvantage of the technique might be that it takes some
time before full production can be realized. In normal mines ex-
panding from a single entry point, many production crews can be
simultaneously advancing in different directions. In this alternative
technique, however, the long main heading must first be completed;
and this generally requires one production unit. Other production
crews could not be efficiently utilized until the main haulage heading
and conveyor belt were completed. During this time, coal production
would also be relatively low. A possible solution to this is develop-
ment of several main headings, fully utilizing capabilities of all
available production crews.
It is also important to note that approximately the same mining al-
ternative technique can be employed in shaft or slope mine to mini-
mize pumping during active operations. The surface elevation of the
shaft or slope must, however, be located at a surface elevation
66
-------
greater than the anticipated head that wilt develop when the abandoned
mine is completely inundated. Down-dip headings can be initially
developed from the shaft or slope to the anticipated limits of mining.
From this point, workings initially expand toward the dip, leaving
low lying mined-out areas to inundate naturally and eliminating much
of the required pumping. Mining can then proceed toward the rise
past the slope or shaft entry to the highest limits of the mine. All
drainage will flow toward the lower end of the workings and, if the
active workings extend up-dip beyond the shaft or slope, in-mine
haulage for that area will be down-grade toward those exit points.
Thus, careful mine planning and slope or shaft placement can effec-
tively reduce both haulage and pumping costs for those mines.
Figures 11 and 12 show the primary advantages of use of one of these
alternative mining techniques as compared to "standard" down-dip
operation.
67
-------
STEP C
LEGEND
Property Line rS-^gg Infiltration
Limit of Mining
Potential Inundation
i-»-H-Working Face (Direction of Production)
NORMAL DOWN-DIP MINING PROCEDURE
Figure II
68
-------
/ \
STEP A
STEP B
STEP C
LEGEND
Property Line
Limit of Mining
Working Face (-^Direction of Production)
Infiltration
Potential Inundation
ALTERNATE DOWN-DIP MINING PROCEDURE
Figure 12
69
-------
SECTION VII
REFERENCES
1. Joy Manufacturing Company. Underground Mining - Introduction
and Development (1972), Volume I.
2. Miller, John T. Seepage and Mine Barrier Width (October 1974),
Fifth Symposium on Coal Mine Drainage Research, Louisville,
Kentucky.
3. Pennsylvania Department of Environmental Resources. Houtzdale
Quadrangle - Northern Half, (1968) Topographic and Geologic
Survey, No. A-85ab.
4. Skelly and Loy. Clearfield Creek and Moshannon Creek, Mine
Drainage Pollution Abatement Project (1973), Pennsylvania
Department of Environmental Resources.
5. Skelly and Loy. Muddy gun Mine Drainage Pollution Abatement
Project (December 1971), Pennsylvania Department of Environ-
mental Resources.
6. United States Department of the Interior. Basic Estimated
Capital Investment and Operation Costs for Undergrgund Bitu-
minous Coal Mines, (1974) Bureau of Mines Information Circular
No. IC-8632.
7. United States Environmental Protection Agency. Processes,
Procedures,, and Methods to Control Pollution from Mining
Activities (October 1973) EPA-430/9-93-011.
8. Water Information Center, Incorporated. Water Atlas of the
United States (1973).
70
-------
SECTION VIII
DEFINITION OF TERMS
Abatement (Mine Drainage Usage) - The lessening of pollution
effects of mine drainage.
Acid Mine Drainage - Any acid water draining or flowing on, or
having drained or flowed from, any area affected by mining.
Aeration - The act of exposing to the action of air, such as, to
mix or charge with air.
Anticline - An upfold or arch of stratified rock in which beds dip
in opposite directions from a crest or axis.
Butt - A small section of rooms and pillars, generally an extension
of an existing panel, in an underground mine.
Down-dip - In the direction of slope of a non-horizontal coal seam.
Effluent - Any water flowing out of the ground or from an enclosure
to the surface flow network.
Heading - A passage or tunnel within an underground mine.
Hydrology - The science that relates to the water systems of the
earth.
Infiltration - Water entering the ground water system and, subse-
quently, underground mine workings through the land surface.
Leaching - Solution of the soluble fraction of a material by flowing
water.
Load (Water Quality Usage) - The quantity of material (acid, iron,
sulfates, etc.) carried by flowing water in solution - generally ex-
pressed in pounds per day.
71
-------
mg/l - Abbreviation for milligrams per liter, a weight to volume
ratio commonly used in water quality analysis. It expresses the
weight in milligrams of a substance occurring in one liter of liquid.
Neujralization - The process of adding an acid or alkaline material
to waste water to adjust its pH to a neutral position.
Outcrop - The surface exposure of bedrock or strata.
Overburden - Nonsalable material that overlies a mineable mineral.
Oxidation - The removal of electrons from an ion or atom.
Panel - A large block of interconnected rooms and pillars worked
either side of a mine's main heading.
pH - Negative logarithm to the base ten of hydrogen ion activity.
pH 7 is considered neutral. Above 7 is basic, below 7 is acidic.
Refuse - Rock that has high carbon content - usually referring to
dark colored coal mining waste material.
Room - A single block of extracted coal.
Slope - An underground mine entry which slants to a coal seam
from the overlying land surface, not from the coal outcrop.
Sludge - Precipitant or settled material from a wastewater.
Slug - Sudden increase in concentration of stream pollutants result-
ing from heavy rainfall rapidly washing leached pollutants from land
surfaces and underground mines.
Stratigraphy - Science of formation, composition, sequence and
correlation of stratified rocks.
Up-Dip - In the direction of maximum rise of a non-horizontal coal
seam.
72
-------
Table 9
CONVERSION TABLE
ENGLISH TO METRIC
MULTIPLY (ENGLISH UNITS) by- TO OBTAIN (METRIC UNITS)
ENGLISH UNIT ABBREVIATION CONVERSION ABBREVIATION METRIC UNIT
acre ac
acre - feet ac ft
British Thermal
Unit BTU
British Thermal
Unit/pound BTU/lb
cubic feet/minute cfm
cubic feet/second cfs
cubic feet cu ft
cubic feet cu ft
cubic inches cu in
degree Fahrenheit °F
feet ft
gallon gal
gallon/minute gpm
horsepower hp
inches in
inches of mercury in Hg
pounds 1b
million gallons/day mgd
mile ml
pound/square
Inch (gauge) psig
square feet sq ft
square inches sq in
ton (short) ton
yard yd
0.405
1233.5
0.252
ha
cu m
kg cal
0.555
0.028
1.7
0.028
28.32
16.39
0.555(°F-32)*
0.3048
3.785
0.0631
0.7457
2.54
0.03342
0.454
3,785
1.609
kg cal/kg
cu m/min
cu m/min
cu m
1
cu cm
°C
m
1
I/sec
kw
cm
atm
kg
cu m/day
km
(0.06805 psig +1)* atm
0.0929 sq m
6.452 sq cm
0.907 kkg
0.9144 m
hectares
cubic meters
kilogram - calories
kilogram calories/kilogram
cubic meters/minute
cubic meters/minute
cubic meters
liters
cubic centimeters
degree Centigrade
meters
liters
liters/second
kilowatts
centimeters
atmospheres
kilograms
cubic meters/day
kilometer
atmospheres (absolute)
square meters
square centimeters
metric ton (1000 kilograms)
meter
* Actual conversion, not a multiplier
73
-------
TECHNICAL REPORT DATA
(Please read Instructions on the reverse before completing)
1. REPORT NO.
EPA-670/2-75-047
2.
3. RECIPIENT'S ACCESSION'NO.
4. TITLE AND SUBTITLE
Dp-Dip Versus Down-Dip Mining
An Evaluation
5. REPORT DATE
June 1975;
Issuing Date
6. PERFORMING ORGANIZATION CODE
7. AUTHOR(S)
John W. Mentz
Jamison B. Warq
8. PERFORMING ORGANIZATION REPORT NO.
9. PERFORMING ORGANIZATION NAME AND ADDRESS
S KELLY and LOY
Engineers - Consultants
2601 North Front Street
Harrisbura. Pa. 17110
10. PROGRAM ELEMENT NO.
1BB040; ROAP 61AAD; Task 016
11. CONTRACT/GRANT NO.
68-01-0465
12. SPONSORING AGENCY NAME AND AuunESS
National Environmental Research Center
Office of Research and Development
U.S. Environmental Protection Agency
13. TYPE OF REPORT AND PERIOD COVEiRED
Final
14. SPONSORING AGENCY CODE
15. SUPPLEMENTAft Y NOTES
16. ABSTRACT
The report presents detailed results of a feasibility study of
down-dip mining, a technique that appears to offer an alternative to
sealing or permanent treatment of polluted effluents from coal mines
after abandonment. The project included an evaluation of a pair of
nearly identical abandoned underground mines - one developed to
rise, one developed to dip - to confirm the theory that discharge
water quality in down-dip mines is substantially better than that in
up-dip mines. An active mine with units operating up-dip and down-
dip was also evaluated to ascertain economic and engineering
limitations, costs in varying situations, and other major advantages
or disadvantages of each mode of operation. Health and safety and
National water quality and economic impacts of widespread use
versus non-use of the technique were also assessed.
17.
KEY WORDS AND DOCUMENT ANALYSIS
DESCRIPTORS
b.lDENTIFIERS/OPEN ENDED TERMS
c. COS AT I Field/Group
Water quality
Coal mines
Cost comparison
*Water pollution
Mine waters
Ground water
Drainage
Evaluation
Mining methods
13B
18. DISTRIBUTION STATEMENT
Release to Public
19. SECURITY CLASS (ThisReportj
Unclassified
21. NO. OF PAGES
82
20. SECURITY CLASS (This page)
Unclassified
22. PRICE
EPA Form 2220-1 (9-73)
74
#U.S.GOYEMUIEirTnillTIN6 OFFICE: 1975-657-593/5392 Region No. 5-11
------- |