;LEA
ATKP
WATER POLLUTION CONTROL RESEARCH SERIES • 14010 DM0 03/70
New Mine Sealing Techniques
For
Water Pollution Abatement
U.S. DfcjeARTMENT OF THE INTERIOR « FEDERAL WATER QUALITY ADMINISTRATION
-------
NEW MINE SEALING TECHNIQUES
FOR
WATER POLLUTION ABATEMENT
by
Halliburton Company
Duncan, Oklahoma 73553
for the
FEDERAL WATER QUALITY ADMINISTRATION
DEPARTMENT OF THE INTERIOR
Program Number 14010 DM0
FWPCA Contract No. 14-12-453
March, 1970
-------
WATER POLLUTION CONTROL RESEARCH SERIES
The Water Pollution Control Research Reports describe
the results and progress in the control and abatement
of pollution of our Nation's waters. They provide a
central source of information on the research, develop-
ment, and demonstration activities of the Federal Water
Pollution Control Administration, Department of the
Interior, through inhouse research and grants and con-
tracts with Federal, State, and local agencies, re-
search institutions, and industrial organizations.
Water Pollution Control Research Reports will be dis-
tributed to requesters as supplies permit. Requests
should be sent to the Planning and Resources Office,
Office of Research and Development, Federal Water
Quality Administration, Department of the Interior,
Washington, D. C. 20242.
For sale by the Superintendent of Documents, U.S. Government Printing Office, Washington, D.C., 20402 - Price $1.50
-------
FWPCA Review Notice
This report has been reviewed by the Federal
Water Pollution Control Administration and
approved for publication. Approval does not
signify that the contents necessarily reflect
the views and policies of the Federal Water
Pollution Control Administration, nor does
mention of trade names or commercial products
constitute endorsement or recommendation for
use.
-------
ABSTRACT
The purpose of this project was to develop and field test new
concepts for watertight mine seal and bulkhead construction applicable
to abatement of acid mine water pollution. Laboratory research deter-
mined proper materials, equipment and techniques for constructing mine
seals or bulkheads. Field testing was conducted on remedial grouting
techniques for constructing mine seals or bulkheads. Two new processes
were developed. One involved a technique of placing a plug of graded
limestone aggregate in a mine drift or portal to neutralize an acid
mine water discharge until a seal was effected. In the process, pre-
cipitation of iron hydroxide gradually closed the pores. A pneumatic
conveying technique permitted placement of up to 550 pounds per minute
of aggregate into a mine drift. The second process consisted of remotely
constructing a mine seal including rear and front bulkheads of a self-
supporting, quick-setting sodium silicate cement specifically developed
for this application. A filler material of expansive cement was used
between the bulkheads to complete the seal. Field testing in West
Virginia substantiated the feasibility of both processes when two aggre-
gate and two bulkhead type seals were placed in abandoned mines which
had drainage flows up to 58 gallons per minute. Cost of the seals are
reported. This report was submitted in partial fulfillment of Contract
No. 14-12-453 between the Federal Water Pollution Control Administration
and the Halliburton Company.
-------
TABLE OF CONTENTS
Page
Abstract
List of Figures
List of Tables
Introduction
A. Synopsis of Previous Study- ------------------ ]
B. Scope and Purpose of Present Study -------------- 3
Conclusions ----- ____________ _____ 5
Recommendations ---------- "_ _______________ 7
Remedial Procedures - Mine Nos. 14-042A & 40-016
A. Remedial Work - Mine No. 14-042A --------------- 9
B. Remedial Work - Mine No. 40-016 19
Bulkhead Development
A. Improvement of Aggregate Placement- --------------33
B. Quick-Setting Mine Bulkheads 38
Permeable Plug Investigation- ------------------- 63
Sealing of High Flow Mine ______ 83
Acknowledgements- -------------------------1Q7
References- ____________________-|09
Patents 111
Abbreviations ---------------------------113
Appendix- ----------------------------_n5
-------
LIST OF FIGURES
Figure
No. Title Page
1 General Location Map- ------------------- 2
2 Mine Location Map --------_-_--_--_____ n
3 Topographic Map - Mine No. 14-042A- - -- 12
4 View of Fractured Coal in Walls and Roof -
Mine No. 14-042A 13
5 Mine Seal Installation - Mine No. 14-042A 14
6 Front View - Expendable Grout Retainer Seal -
Mine No. 14-042A- 15
7 Fluid Pumping Operation - Mine No. 14-042A -- 16
8 Section Through Mine No. 14-042A 18
9 Mine Map - Mine No. 40-016 22
10 Aggregate Plug Detail - Mine 40-016 23
11 Leaking Grouted Aggregate Seal - Mine No. 40-016- - - 24
12 Grout Hole Location - Mine No. 40-016 25
13 Auger in Drilling Position - Mine No. 40-016 26
14 Special Grout Hole Packer - Mine No. 40-016 26
15 Bottom Discharge Manifold - Pressure Tank _____ 35
16 Pneumatic Conveying Test - 3/4-Inch Aggregate ------- 36
17 Buildup of Gypsum Cement Slurry in Yard Test- ------- 43
18 Nozzle Development Test 43
19 Special Nozzle for High Viscosity Slurries 44
20 Slurry Distribution in Bulkhead Construction -
Simulated Mine Test 45
21 Completed Bulkhead - Simulated Mine Test 46
22 Mine Map - Mine No. 62-008 48
-------
LIST OF FIGURES - Continued
Figure
No. Title Page
23 Main Portal - Mine No. 62-008 Before Cleanup 49
24 Work Site After Cleanup - Mine No. 62-008 49
25 Inside of Main Portal Before Cleanup - Mine No. 62-008 50
26 Beginning of Bulkhead Construction in Mine No. 62-008 52
27 Completed Bulkhead in Mine No. 62-008 52
28 Section - Quick Setting Bulkhead - Opening No. 4 -
Mine No. 62-008 53
29 Mixer Nipple 54
30 Yard Tests of Slurry Using Heated Water ---------- 56
31 Remote Operation of Equipment for Placement of Slurry - - - 59
32 Grout Pipe Location - Opening No. 5 - Mine No. 62-008 60
33 Section - Mine Seal - Opening No. 5 - Mine No. 62-008 61
34 Remedial Work - Mine No. 62-008 62
35 Laboratory Apparatus for Permeability Studies ------- 65
36 Completed Trough - Permeable Plug Tests -- ____- 69
37 Trough - Inlet Compartment and Valves ----------- 69
38 Flow Through Limestone in Wooden Trough - Permeable
Plug Test No. 1 70
39 Water Level in Limestone Filled Trough - Permeable
Plug Test No. 1 71
40 Plugging Profile - Wooden Trough - Test No. 1 ------- 72
41 Section - Compartments - Wooden Trough Tests- ------- 74
42 Trough - Overall View - Test No. 2---- ______ 75
43 Trough - Discharge End - Test No. 2- - ______ 75
44 Trough - Input Arrangement - Test No. 2- _--- 75
-------
LIST OF FIGURES - Concluded
Figure
No. Ti tle Page
45 Placing Aggregate in Simulated Mine Openings- ------- 80
46 Drift Mine for Permeable Plug Installation 80
47 Mine Interior - Permeable Plug Site ----- — -____ 81
48 Placing Limestone Seal in Mine Drift- --- _.__ 81
49 Location Map - Mine No. RT5-2 84
50 Mine Map - Coalton No. 3 Mine (RT5-2) 86
51 View of Mine No. RT5-2 87
52 Plan View of Remedial Construction - Mine No. RT5-2 88
53 Grout Pipe Arrangement - Seal No. 1 - Mine No. RT5-2- - - - 92
54 Beginning of Front Bulkhead Construction- ----- — -- 93
55 Completed Front Bulkhead - Mine No. RT5-2 94
56 Construction Detail - Seal No. 1 - Mine No. RT5-2 95
57 Opening Excavated Right of Main Portal - Mine No. RT5-2 - - 99
58 Opening No. 2 After Exposing Front - Mine No. RT5-2 100
59 Dust Created with Pneumatic Conveying of Aggregate
Containing Agricultural Lime --101
60 Grout Pipe Placement - Opening No. 2 - - - - 102
61 Section View of Seal - Opening No. 2-- 105
62 Completed Installation - Opening No. 2 106
63 Geologic Section - Mine No. 14-042A 117
64 Geologic Section - Mine No. 40-016 117
65 Geologic Section - Mine No. 62-008 118
66 Geologic Section - Mine No. RT5-2 118
-------
LIST OF TABLES
Table
No. Title Page
1 Fluid Property Test - Mine No. 14-042A 119
2 Mine Fluid Data - Mine No. 14-042A 120
3 Water Monitoring Data - Mine No. 14-042A- - 121
4 Water Analyses - Mine No. 14-042A 122
5 Water Monitoring Data - Mine No. 40-016 -124
6 Water Analyses - Mine No. 40-016 - Opening No. 1 127
7 Water Analyses - Mine No. 40-016 - Opening No. 2 129
8 Pneumatic Conveying Test - 3/4-Inch Aggregate ------- 131
9 Organic Foam Materials Investigation- ----------- 132
10 Inorganic Foam Tests- -------------------133
11 Initial Test Data - Slurry Composition- ---------- 134
12 Additional Test Data - Selected Slurries 138
13 Quick-Setting Slurry Compositions -139
14 Data for Selected Silicate Cement Slurry ----140
15 Low Temperature Tests of Bentonite-Sodium Silicate
Slurry 141
16 Water Monitoring Data - Mine No. 62-008
Specified Openings- -------------------142
17 Water Analyses - Mine No. 62-008 145
18 Screen Analysis - Greer Limestone Samples
Screen Analysis - Harrold Limestone Samples -------- 146
19 Retention Time Flow Test - Harrold #12 Limestone- ----- 147
20 Flow Test Data - Modified Greer #13 Graded Limestone - - - 148
21 Flow Test Data - Modified Greer #13 Limestone
Containing 1% BaCOs 148
-------
LIST OF TABLES - Concluded
Table
No. Title Page
22 Fluid Flow Data - Permeable Plug Test No. 1 149
23 Water Sample Analyses - Permeable Plug Test No. 1 -
Mine No. 40-085 - 150
24 Fluid Flow Data - Permeable Plug Test No. 2 - 152
25 Fluid Flow Data - Modified Permeable Plug Test No. 2 153
26 Water Sample Analyses - Permeable Plug Test No. 2 154
27 Fluid Data and Water Analyses - Permeable Plug
Test No. 3 159
28 Screen Analyses - Limestone Used in Mine No. RT5-2 - - 160
29 Monitoring Data - Seal No. 1 - Mine No. RT5-2 161
30 Monitoring Data - Seal No. 2 - Mine No. RT5-2 162
31 Water Analyses - Mine No. RT5-2 163
-------
INTRODUCTION
Synopsis of Previous Study
A previous contract to the Halliburton Company consisted of a
feasibility study in four parts on the application of various grouting
agents, techniques and methods to the abatement of mine drainage pollution.
This study, conducted from December 1966 to May 1968, included a survey
of mine drainage pollution in the Upper West Fork River Sub-basin near
Clarksburg, West Virginia. This area is located on the map attached as
Figure 1.
The Monongahela River is formed by the confluence of the West
Fork and Tygart Valley Rivers, which occurs approximately five miles
north of the northern boundary of Harrison County, near Fairmont, West
Virginia. The Halliburton study was concerned primarily with that por-
tion of the West Fork River Basin which terminates immediately south of
Clarksburg in Harrison County and extends to the headwaters in Upshur
County. The Sub-basin contains approximately 400 square miles and more
than 200 drift mines.
Clarksburg, the fifth largest city in the State and seat of
Harrison County, is the major center of population and industry in the
Upper West Fork River Sub-basin. Situated at the base of the Sub-basin,
the city obtains its water supply directly from the West Fork River.
Numerous other communities are interspersed throughout the Sub-basin and
also take their water supplies from the river or its tributaries.
The feasibility study included an initial survey made by
reconnaissance parties throughout the Upper West Fork Sub-basin. All
mined areas as revealed from information furnished by the Monongahela
River Mine Drainage Remedial Project were included in the survey.
The reconnaissance party located and identified the openings -
which were evident within the Sub-basin and drainage water was sampled
and analyzed to determine the acid and mineral content. The geology of
the mine sites was also recorded. A total of 228 drift mines or open-
ings were discovered in this survey. A resurvey was then conducted on
60 mines which evidenced measurable flows during the preliminary survey,
or had drainage with high mineral content.
Possible remedial techniques to abate the flow of mine drain-
age from the mines in this area were studied and a listing of the various
methods was made. A drift mine which had an acidic drainage coming from
the mine opening was selected for the installation of a mine seal. The
seal installed consisted of four layers of cloth retainers which were
progressively filled with a quick-setting cement grout. Each layer was
allowed to conform to the shape of the mine floor, walls or roof as the
material hardened. When complete, the stack of retainers formed a seal
which conformed to the shape of the opening.
-1-
-------
CLEVELAND
PENNSYLVANIA
PITTSBURGH
MARYLAND
MONONGAHELA
RIVER BA91N
CREEK BASIN
UPPER WEST FORK
RIVER SUB-BASIN
^CHARLESTON
VIRGINIA
WEST
VIRGIN IA
FIGURE I - GENERAL LOCATION MAP
-2-
-------
Thjree other mine complexes were selected which were felt to be
representative of the flow and acid drainage of those within the Basin.
Complete plans and specifications were drawn up for abatement construction
work on the mine openings, utilizing the grout retainer seal technique as
well as other sealing methods.
Additional work in the study included research on the grout re-
tainer technique of sealing a mine, with emphasis being given to develop-
ment of a cement slurry which would permit the seal to be placed in one
working day. This would reduce the cost which had been incurred on the
one already placed, since that application took several days. A similar
type retainer was developed for use in 40-inch auger holes. Several seals
were successfully placed in auger holes in field trials. Two mine drifts
were sealed remotely by the technique of placing aggregate in the mine
pneumatically and subsequently grouting this aggregate to form a 25-foot
plug in each mine drift opening.
Scope and Purpose of Present Study
The work reported herein is a sequel to the previous feasibility
study. The first phase of the work involved remedial procedures to correct
leakage which was issuing from the mines sealed in the previous contract.
The seal installed in Mine No. 14-042A was not leaking, but there was water
issuing from a subsided area at the left front of the mine seal at a rate
of about one-half gallon per minute. The two openings of Mine No. 40-016
had previously been sealed by placing aggregate plugs in the mine and
grouting the aggregate. Water was leaking from these openings at the rate
of about 3 gallons per minute from Opening No. 1 and 4.5 gallons per minute
from Opening No. 2.
Another phase of the present study was a research investigation
to improve the pneumatic method for placement of aggregate into a mine
drift and thereby make the process more economically feasible.
Additional research was conducted to determine other techniques
of seal installation and construction which might provide an economic or
practical advantage over the grout retainers used in the mine seal con-
structed in the previous study. The unsuitability of the grout retainer
seal in mine drifts containing center posts was another consideration
for further research in mine seals.
Research was also conducted to determine the feasibility of
placing a permeable plug within a mine drift. A permeable plug is a
partial seal which allows the escape of some mine water through the plug
and at the same time provides in-place treatment for the fluid as it
passes -through the plug.
A high-flow mine was selected for the field application of the
techniques developed in the other phases.
-3-
-------
CONCLUSIONS
Numerous findings have resulted from the tests and evaluations
performed under this contract. The following conclusions are drawn from
these findings:
Bulkhead Development
(1) A system was successfully developed to move aggregate
pneumatically at rates up to 550 pounds per minute with a minimum amount
of downtime. The method utilized two alternately loaded and discharged
pneumatic tanks. This could approach a continuous operation by increas-
ing the size of the tanks and/or the number of tanks.
(2) It was found that pneumatic handling of aggregate which
contains fine or powdered limestone and any appreciable amount of mois-
ture tends to clog the conveying pipe. Every effort should be made to
keep the material dry while loading or conveying.
(3) To convey aggregate pneumatically from pressurized tanks,
the use of a special air manifold as described in the report is mandatory.
(4) Investigation of two organic foam materials led to the
conclusion that neither would be practical for building bulkheads in
drift mine applications. Equipment and material costs, together with
knowledge of uncontrollable mine atmospheric conditions which could
adversely affect foam density, were the basis for elimination of poly-
urethane foam. Urea-formaldehyde foam was concluded as unsatisfactory
because of low strength, low bulk density, high equipment costs and poor
resistance to biological, dilute acid or water attack.
(5) Laboratory investigation of inorganic foam materials for
possible use in building temporary drift mine bulkheads indicated that
these materials were fragile with poor water resistance. It was con-
cluded that these characteristics would make this material unsatisfactory
for bulkheads in mine drifts.
(6) A sodium silicate cement slurry with quick-setting and
self-supporting properties was developed. This material was satisfac-
tory for use in remotely building bulkheads in mine drifts.
(7) Special equipment for mixing and placing the quick-setting
slurry was developed. Techniques using this equipment and the special
slurry were proven to be satisfactory. It is concluded that this method
is a practical and applicable method for construction of mine seals.
(8) Based on the results of field tests, it was concluded that
a mine seal, composed of rear and front bulkheads of a quick-setting
slurry with expansive type cement filler between, would be satisfactory
to seal a mine drift to hold a required head of impounded mine water.
-5-
-------
(9) Obtaining access rights for field applications is difficult
and can require a considerable expenditure of time and money.
Permeable Plug Investigation
(10) Laboratory investigation determined that a graded limestone
aggregate could be used as a plug to neutralize acid mine water passing
through the plug as it gradually seals the pore space due to precipitation
of iron hydroxide. The initial success on a field application of a plug
of graded aggregate in a mine drift tends to confirm this laboratory find-
ing. Based on these findings, this method appears promising as a practical
approach to acid mine water impoundment and in-place treatment.
Sealing of a High Flow Mine
(11) Field tests were conducted to determine the feasibility of
placing a mine seal in a high flow mine. Based on findings from these tests,
it is concluded that a mine seal can be constructed in a high flow mine
using the techniques developed under this contract.
(12) It was concluded that an expansive cement filler is preferred
to a grouted aggregate filler for double bulkhead type mine seals when ex-
perience to date in constructing mine seals, ease of placement, economic
factors and possible need of remedial work are considered.
(13) The use of agricultural lime blended in the graded aggre-
gate caused dust problems. The use of a water spray at the discharge helped
some but did not solve the problem. It was concluded that this problem must
be solved before powdery material can be used extensively.
Remedial Procedures - Mines No. 14-042A and 40-016
(14) The injection of gelled fluid into a previously sealed mine
was not successful in reducing leakage. No evidence of gelled fluid return
was observed at the point of leakage on Mine No. 14-042A. It was concluded
that the gelled fluid did not communicate with the leakage point. Such
communication is essential in order to seal with this type of treatment.
(15) Remedial grouting with cementitious material was conducted
unsuccessfully on an aggregate seal which was placed and grouted in Mine No.
40-016 under a previous contract. It is apparent that it is very difficult
to consolidate aggregate in place and obtain a satisfactory seal due to the
necessity of containing the grout within the aggregate. Although many of
the fluid channels were apparently closed during the remedial grouting,
enough additional channels remained open to permit the flow to continue un-
abated. Because of the rate of leakage through these two grouted aggregate
seals which were emplaced during the previous contract and the lack of success
in sealing the leaks, it is concluded that the grouted aggregate type seal
is not practical.
-6-
-------
RECOMMENDATIONS
Based on the conclusions drawn In this study, the following
recommendations are submitted:
(1) When conditions permit, it is recommended that drift open-
ings of abandoned mines be sealed utilizing rear and front bulkheads of
self-supporting cement and a light, expansive type cement between the
bulkheads.
(2) It is recommended that methods other than placement of
aggregate in drift openings with subsequent grouting be used as a mine
seal.
(3) It is recommended that further development work be under-
taken to provide more efficient placement equipment for the construction
of self-supporting bulkheads.
(4) It is recommended that research be conducted to develop
the equipment and techniques for constructing quick-setting bulkheads in
inaccessible mine drifts through vertical bore holes.
(5) It is recommended that further research be conducted to
determine the optimum materials for use in constructing a permeable
plug type mine seal.
(6) It is recommended that further research include a study
to determine a suitable means of eliminating the extreme dust problem
encountered when pneumatically conveying dusty material into mine
openings.
(7) It is recommended that efforts to obtain access rights
to property be initiated well in advance of any proposed field work.
-7-
-------
REMEDIAL PROCEDURES - MINES NO. 14-042A & 40-016
Two mines, sealed during a previous contract, were selected for
remedial work. Mine No. 14-042A had been closed by the use of four ex-
pendable grout retainers. Expendable grout retainers were fabricated
cloth containers of nylon or cotton, varying in length from 20 feet for
the bottom retainer to 10 feet for the top. Each retainer was succes-
sively filled with cement grout slurry to conform to the shape of the
drift opening. Leakage at the rate of 1/2 gallon per minute was flowing
from a fractured coal surface to the 1 eft of the seal. The work was
performed in an attempt to stop or significantly reduce this leakage.
Two drift openings were exposed at Mine No. 40-016. These
openings had been sealed by aggregate plugs 25 feet in length which were
placed in the mine pneumatically and grouted. Leakage was observed under
the base and across the front width of each plug. Opening No. 1 was flow-
ing about 3.0 gallons per minute and 4.5 gallons per minute was leaking
from Opening No. 2. The remedial grouting was performed in an attempt
to reduce or stop the leakage.
A map showing the location of these two mines is attached as
Figure 2.
Remedial Work - Mine No. 14-042A
This is a small isolated mine of approximately 5 acres located
about one mile south of Clarksburg, West Virginia and east of the West
Fork River and State Secondary Route 25. The Pittsburgh coal seam has
been extensively drift mined and partially strip mined, leaving negli-
gible coal reserves. The coal dips steeply to the east.
A flow of 18 gallons per minute from the mine was measured
just prior to sealing the drift opening. The flow discharges into Doll
Run, a tributary to the West Fork River. The mine water was found to
have a pH of 2.6, an acidity of 2750 mg/1 in terms of CaCOs and an iron
content of 558 mg/1. The acid production rate was 616 pounds per day.
Mining in this area began in the early part of the 20th Century.
Most of the mines were closed in the 1930's or prior to World War II. An
extensive search was made for a map of this mine, but one could not be
located. A topographic map was obtained and is included as Figure 3. A
geologic section of the mine is included (Figure 63 - Appendix).
This mine was selected for application of a mine seal because
of its remote location, small but easily accessible area and highly
acidic discharge. The use of expendable grout retainers for a seal in-
stallation was selected on the basis of research tests. This type seal
consisted of successive layers of expendable grout retainers filled
with a cement grout slurry. An expendable grout retainer was a rec-
tangular nylon or cotton cloth container which varied in size from
20 x 10 x 3 feet to 10 x 8 x 3 feet.
-9-
-------
As the seal was installed, the 20-foot long nylon retainer was
placed on the floor of the mine and inflated with cement slurry to con-
form to the shape of the opening. After the first retainer was suffi-
ciently hard to withstand a load of about 3 psi, a second nylon re-
tainer which was 16 x 10 x 3 feet was placed on the first. Figure 4
shows this retainer in place against the walls of the mine. The process
was repeated using a third nylon retainer which was 14 x 10 x 3 feet.
The 10-foot cotton retainer was used to obtain sufficient height to com-
pletely fill and seal the opening. A drawing of this seal showing the
installation and mine configuration is attached as Figure 5. It should
be noted that the roof and walls of the mine were of coal and the floor
was slate. The coal was irregular in shape and contained many fractures,
as can be clearly seen in Figure 4. A detailed account of this instal-
lation is given in the Halliburton Company report, Part II - "Selection
and Recommendation of Twenty Mine Sites", August 23, 1967, p. 137.
After the mine seal was installed, the cumulative leakage from
small leaks in the coal surfaces around the seal was measured at 1.5 gallons
per minute, a reduction of 92% from the leakage measured from this mine be-
fore the seal installation.
Preparation was made to reduce this leakage by grouting around
the lower grout retainer using Halliburton PWG grout fluid. This fluid
was an acrylamide monomeric solution which had a 10-second set time. An
injection of 20 gallons of grout reduced the flow to 0.75 gallons per
minute. Further grouting around the second retainer using 30 gallons of
grout reduced the leakage to 0.33 gallons per minute. It then appeared
that the remaining leakage was coming through coal fractures several feet
to the left of the seal face, so grouting was discontinued. Figure 6
shows a view of this mine seal at the time of the present contract initi-
ation.
The plan for remedial work on the present contract was to grout
the leakage points with the same grout fluid used previously. Before
this work was initiated, the approach was changed. Accordingly, the State-
ment of Work was modified to conduct remedial measures by injecting a low
viscosity polymer or gelling agent into the mine behind the seal to attempt
to seal the points of leakage by flowing the material into the leakage
points from the inside of the mine.
A gel material made using Karaya gum was considered, but was
rejected when it was found that the price of $0.415 per pound for this
material would result in a high treatment cost. An alternate material
selected was a gel material consisting of bentonite, shredded cane fiber
and water. The cost for the dry materials was $0.028 per pound. The
cost for the gelled fluid after placement in the mine was $0.069 per
gallon.
Prior to beginning remedial operations, an initial measurement
was taken of the water level of the impounded water. The measurement
-10-
-------
OOODHOPE
LEWIS COUNTY
COUNTY
FIGURE 2 - MINE LOCATION MAP
-11-
-------
HENRY BASSEL - SURFACE
VA. ALLEN CORP. - COAL
COAL CONTOUR INTERVAL - IFT
SURFACE CONTOUR INTERVAL -5FT
I2FTCOVER LINE -—-* — '
:OAL OUTCROP LINE
BUELLA FIDLER-SURFACE
VA.ALLEN CORP- COAL
NOTE
ASSUMED ELEVATION
100 150 200
SCALE, FT.
MINE SEAL LOCATION
T7
FIGURE 3-TOPOGRAPHIC MAP-MINE NO I4-042A
-12-
-------
FIGURE 4 - VIEW OF FRACTURED COAL IN WALLS AND ROOF - MINE NO. 14-042A
-------
3" MAIN PIPE
I" GROUT PIPE
I" 01*. PLASTIC
SHOUT PIPE
I" DIA. PLASTIC
OBOUT PIPE
20'MTUW
RETAINER
I'fiROUT PIPC
j" Hum »
SAMPLING TUK
FIGURE 5
MINE SEAL INSTALLATION
MINE NO.I4-042A
HARRISON COUNTY, WEST VIRGINIA
-------
FIGURE 6 - FRONT VIEW-EXPENDABLE GROUT RETAINER SEAL-MINE NO. 14-042A
-15-
-------
LEAKAGE THROUGH
FRACTURED COAL
SURFACES
OBS. WELL
250 BBL.
WATER TANK
BOTTOM DRAIN
VALVE
GROUT RETAINER SEAL
IN DRIFT OPENING
FIGURE 7- FLUID PUMPING OPERATION-MINE NO. I4-042A
-------
showed a depth of 5.0 feet of fluid. The leakage rate was measured at
0.4 gallon per minute.
Equipment utilized for the remedial work is shown in the sche-
matic drawing of the Fluid Pumping Operation attached as Figure 7. This
drawing also designates the points of leakage.
The work was initiated by mixing a small batch of gelled fluid
consisting of 5 sacks (500 pounds) of Wyoming bentonite and 800 gallons
of fresh water. Samples were taken at intervals throughout the job to
determine viscosity and filter loss using the standard API testing pro-
cedure as given in American Petroleum Institute publication RP 13B,
November 1962. A table listing all fluid properties obtained in these
tests is included in the Appendix as Table 1.
The fluid was pumped into the mine through the 3-inch drain
line which had previously been installed as a part of the mine seal. As
the fluid was being pumped into the mine, a fiberous material of shredded
cane fiber (Fibertex) was added to act as a sealant for any fractures
leaking fluid. The rise in the mine fluid level after the addition of
the first batch of gelled fluid was 1-1/4 inches.
After the pilot batch was pumped into the mine, it was decided
that four larger treatments would be used for a total of about 40,000
gallons of gelled fluid. Injection of one batch per day was selected.
This would allow time for observation and evaluation of each treatment
separately.
A batch of 9,500 gallons of gelled fluid was then mixed, using
7,000 pounds of bentonite and 9,200 gallons of fresh water, and pumped
into the mine through a 3-inch line. Fibertex was added continuously at
the rate of 1/2 pound per 100 pounds of bentonite. This was increased
to one pound per 100 pounds after 2,500 gallons had been pumped into
the mine. This ratio was then retained for the remainder of the remedial
treatment.
Three additional batches were mixed using 10,000 gallons of
fresh water and 7,500 pounds of bentonite. These batches, each consist-
ing of 10,300 gallons of gelled fluid and 75 pounds of Fibertex, were
pumped into the mine on the next three consecutive days after the first
treatment.
During the final pumping stages, some of the viscous fluid was
observed seeping through the coal seam on the upper right side of the
top retainer. Leakage had not been observed in this area previously.
No further leakage was noted after the treatment.
A total of 41,200 gallons of gelled fluid was pumped into the
mine, utilizing 30,000 pounds of Wyoming bentonite and 295 pounds of
shredded cane fiber blend. A schematic drawing of the mine is attached
as Figure 8 to show the condition following the remedial work.
-17-
-------
-RECORD™ HOUSING
MATER LEVEL RECORDER
/- 7* SCH 40 PIPE CASIM
IRRE6ULAR ROOF COAL
/-APPROX.ELEV. 5.0*. y^TT^
EXPENDABLE GROUT
RETAINERS —»
^v^ifx^J/i;^^^!^^
5 DRAIN PIPE.
6ELLED FLUID
„- PUT INTO MINE
:?$ THROU6H THIS
LINE
FIGURE 8 - SECTION THROUGH MINE NO. 14-042 A
-------
The flow rate at the completion of the treatments was 0.55
gallon per minute, an increase of 0.15 gallon per minute over the ini-
tial rate. The leakage rate has continued to average about 0.5 gallon
per minute since treatment. This is approximately the same as before
the treatment, indicating that the treatment did not accomplish the
desired results.
The gelled fluid apparently did not fill all the void space
behind the mine seal. The rise in fluid level directly behind the seal
without any increase in leakage seemed to indicate that the viscous fluid
did not travel into the left side of the mine from which the leakage was
discharging. A roof fall in that area could have restricted the movement
of the more viscous fluid. It is also possible that the water leaking
through the coal fractures to the left of the seal might be coming di-
rectly from water percolation and not from within the mine. This might
also explain the change of leakage rate without any change in water level
in the mine. The observation of some gelled fluid discharging from an-
other point during the treatment seems to substantiate the theory that
the fluid would discharge through a fracture if the fluid had access to
it from within the mine.
The water table within the mine and the discharge rate actually
rose about 20% during this time (Table 2 - Appendix). A detailed tabu-
lation of additional monitoring on this mine is included in the Appendix
as Table 3 which indicated that this remedial work was not successful in
controlling the discharge. Water analyses data are found in Table 4 of
the Appendix. It can be seen that the mine water was initially high in
acid, iron, sulfate and aluminum content but appeared to be decreasing
as time progressed.
The cost for the materials and equipment used in this remedial
work on Mine No. 14-042A was $2,771.00. Access rights cost an additional
$300.00 and site restoration required an expenditure of $279.00.
Remedial Work - Mine No. 40-016
This is a small, abandoned commercial drift mine located south
of Clarksburg, West Virginia, midway between West Milford and Lost Creek
in Harrison County. Drift openings can be observed in both the Pitts-
burgh and Redstone coal seams with the major headways located in the
Pittsburgh seam. Both Pittsburgh and Redstone coal have been strip
mined.
Mining was begun in this area around 1904 and continued until
about 1930. The mine was abandoned but still had some coal reserves
present. In addition to the drift and strip mining, extensive augering
has been performed in the Redstone coal seam on the south side of the
mine island. Any augering which may have been done in the Pittsburgh
coal seam is obscured by land reclamation backfill. No subsidence was
observed in the overburden above the mine. A geologic section of the
-19-
-------
mine is Included as Figure 65 to show the composition of the overburden.
Drainage from this site discharges into Stone Pot Run and thence
to Lost Creek. Flows from 7 to 150 gallons per minute have been measured.
The mine water had a pH of 2.9 to 3.4, an acidity from 306 to 396 mg/1 in
terms of CaC03 and iron varying from 120 to 190 mg/1.
Two main openings from the Pittsburgh coal were found. These
openings are shown on the mine map attached as Figure 9 and designated as
Opening No. 1 and No. 2.
In the previous study, Openings No. 1 and 2 were selected for
installation of a mine seal. An expendable grout retainer seal was con-
sidered for this application, but was rejected when a row of center posts
were discovered in Opening No. 2. Based on results of research work, it
was decided that limestone aggregate could be placed in the drift opening
pneumatically and then consolidated into an impervious mass with a cement
grout. This method was used for sealing the two openings of Mine No. 40-016.
Each opening was determined to be 12 feet wide and 7 feet high.
Opening No. 1, the main entrance, had a jagged and irregular wall but had
no roof falls and was supporting itself well with a minimum of timbering.
Opening No. 2, the fanway, had a very rough, irregular floor and roof and
a row of center posts. The roof had collapsed and was being supported
solely by the timbers for 12 to 15 feet from the opening.
Prior to construction of the seal, the opening was enlarged to
full drift size by removing backfill and sloughed highwall. The floor,
walls and roof were cleaned by jetting with water with a pressure of
150 psi. This cleaning was performed remotely using a maneuverable jet
developed in the previous contract.
To construct the seal, 3/4- to 1-1/2-inch graded limestone was
placed in the mine pneumatically. The placement was accomplished remotely
using the placement dolly developed in the previous contract. The rock was
loaded in pneumatic tanks and moved with air at 25 psi pressure through
4-inch hose to the aggregate placement dolly. The aggregate guide plate
on the 4-inch pipe of the placement dolly permitted the operator to re-
motely control the aggregate placement by rotation of the pipe or movement
of the dolly.
The center portion of the aggregate seal contained grout pipes
which were installed before placement of the limestone. A layer of minus
3/8-inch aggregate was put on either side of the center section to help
contain the grout. The construction details can be noted in the cross
section of the installation which is attached as Figure 10. A detailed
account of the construction project is given in the Halliburton Company
report, Part IV - "Additional Laboratory and Field Tests for Evaluating
and Improving Methods for Abating Mine Drainage Pollution", May 21, 1968,
p. 126.
-20-
-------
After the seals were installed, leakage was noted coming around
and through the seals and was also observed across the width of the front
base. However, it was reduced to about 7 gallons per minute from both
openings, a reduction of 85% from the average of four previous measure-
ments.
Under the present contract, it was specified that remedial
grouting would be performed to attempt to stop or reduce the leakage
from the two openings. The main body of the limestone seal was to be
grouted using a cement type grout.
The grout pipes which were extending from the existing aggre-
gate seal were uncovered as shown in Figure 11. An effort was made to
pump into these pipes to determine if remedial grouting could be accom-
plished in this manner. It was discovered that the grouting in the
previous study had left the pipes plugged so that it was impossible to
pump into any of the pipes. It was therefore decided that additional
grout holes would be drilled at the face of the highwall on each side
of the opening at an angle extending to a point about midway along the
length of the aggregate. A crew was retained to survey the site and
stake the location for the auger holes into the highwall. Four hole
locations were marked and designated for drilling, with two holes into
each seal. A plat showing the location of each of these holes as deter-
mined by the survey is attached as Figure 12.
Stakes were-placed on the base line and a steel pin positioned
at the proposed point of entry. These were used to align the hole so
that the augering machine could then drill the grout hole at the correct
angle. Figure 13 shows the augering machine mounted on a flatbed trailer
preparing to drill the No. 1 grout hole. A 6-inch diameter hole was
drilled at a height of 84 inches above the base of the coal. Drilling
was continued until a depth of 33-1/2 feet had been drilled. At this
point, a 2-foot void was encountered and the hole began to seep water.
Drilling was continued for an additional 3 feet. At this time, the hole
began to produce a considerable amount of water, indicating the possi-
bility that a channel had been encountered leading to the rear of the
seal.
A 2-inch pipe was used in the drilled hole as a grout pipe.
A special packer was made to use in grouting the 2-inch pipe in the
6-inch drilled hole. This packer was made by wrapping brattice cloth
around the Tower part of the 2-inch joint of pipe and banding it on the
inward end. The length of the finished packer was 2 feet. This packer
is shown in Figure 14.
The 25-foot joint of 2-inch pipe was placed into the grout
hole to its full depth, then pulled back approximately 1-1/2 feet to
tighten the packer and shut off the water flow around the pipe. A 25-
foot joint of 1-inch pipe was then put into the hole alongside the
2-inch pipe until it was at a point about 6 inches from the set packer.
-21-
-------
AUGEREO REDSTONE
PITTSBURGH SEAM
C.V. SMITH
xAUGERED REDSTONE
L/8 PITTSBURGH SEAM
V
I
I
f
l-_
h
•
OMA STOUT
s
i
J^AUGCRED REDSTONE
8 PITTSBWRGM SEAM
MINE NO.40-016
FIGURE 9
-22-
-------
3/4." TO i '/2 "AGGREGATE
.GROUTED SECTION
no
LO
i
SECTION THRU OPENING NO. 2
"TO \yz
AGGREGATE
I GROUT PIPES
l" GROUT PIPES
2" DRAIN LINE
t '+
1
,£'4-
A GROUT -^
LPIPES
^ ,2" DRAIN
f (LINE .
.
4^-60'±
m
A GROUT —
^PIPES
^ ,2" DRAIN
f (UNE ,
v-— ^5
m
i
\
OPENING NO. 2-FAN WAY
OPENING NO. I - ENTRANCE
FIGURE 10-AGGREGATE PLUG DETAIL-MINE 40-016
-------
FIGURE 11 - LEAKING GROUTED AGGREGATE SEAL - MINE NO. 40-016
-------
I
ro
en
i
10 20
SCALE, FEET
40
46°
| BASE LINE "C"
FIGURE 12-GROUT HOLE LOCATION -MINE NO. 40-016
-------
FIGURE 13 - AUGER IN DRILLING POSITION - MINE NO. 40-016
FIGURE 14 - SPECIAL GROUT HOLE PACKER - MINE NO. 40-016
-26-
-------
This pipe was used in the grouting of the 2-inch pipe in the hole. A
valve was placed on the 2-inch line and closed. This stopped the flow
of water from the mine through this pipe.
In like manner, grout hole No. 2 was drilled at a point 56
inches above the base of the coal. The depth of this hole was 45 feet
before the hole started discharging water. A 25-foot joint of 2-inch
pipe with a cloth packer was placed in the hole in a similar manner to
grout hole No. 1. The packer was set and the 1-inch pipe placed along-
side the 2-inch pipe for grouting purposes. Grout hole No. 3 was drilled
52 inches above the base of the Pittsburgh coal. Water was encountered
at 30 feet, but the hole was drilled to a depth of 42 feet to provide
entry into the aggregate plug. The packer was placed on the pipe and
set in the hole and the 1-inch pipe was placed alongside for grouting,
as was done on the previous holes. The No. 4 grout hole was drilled
57 inches above the base of the Pittsburgh coal. As can be noted on
Figure 12, the depth required for this hole is much less than the other
holes. Water and limestone chips were encountered at a depth of 23 feet.
Drilling was discontinued at a depth of 26 feet. The pipe containing
the packer was placed in the hole with the accompanying 1-inch pipe.
The grouting of the 2-inch pipe inside the 6-inch drilled hole
was necessary to permit grouting through the pipe. This was accomplished
by closing the hole entrance with clay to enclose the 2-inch pipe and a
short length of 1-inch pipe at the top of the hole. As cement slurry was
pumped into the hole through the long 1-inch pipe, the hole filled com-
pletely around the 2-inch pipe until the grout flowed out the short
1-inch top pipe. About 6 sacks of cement were required for each hole to
successfully grout the 2-inch pipe.
After allowing sufficient time for the cement grout to harden,
the grouting installations were tested using fluorescein dye. In grout
hole No. 1, 800 gallons of fluid were injected through the 2-inch pipe
at the rate of 20 gallons per minute with 20 psi pressure for the first
400 gallons, and 40 gallons per minute at 100 psi pressure for the last
400 gallons. The return of dye was not evident in the water being mon-
itored from the mine. In grout hole No. 2, 125 gallons of fluid contain-
ing dye were injected in 6 minutes. The dye water was observed coming
through the coal on the left-hand side of the opening about the middle of
the coal vein, extending from the front edge of the aggregate to about
8 feet in front. This seemed to indicate the vertical and horizontal
communication that would be prevalent in the proposed remedial work. In
grout hole No. 3, 125 gallons of dye water were injected in 8 minutes.
The dye water was observed flowing from Opening No. 2 through cavings
and debris on the right side and across the top of the aggregate. In
grout hole No. 4, 250 gallons of dye water were injected over a 14-minute
period. The dye was observed coming across the top of the aggregate seal,
covering about three-fourths of the width. Satisfactory returns were
obtained from grout holes 2, 3 and 4, indicating that the grout fluid
would probably move through the aggregate to the front of the aggregate
-27-
-------
at the highwall. Grout hole No. 1 did not show any returns, so a grout
fluid injected into this hole may not go to the right area to seal off
any leakage.
A grout slurry was selected and approved consisting of pre-
blended 50% portland cement and 50% fly ash by volume, with 18% salt
and 0.4% Halad-9. Halad-9 is a Halliburton Services additive to prevent
water from being lost into permeable formations due to applied pressure.
Based on Halliburton's many years of grouting experience, this particular
grout slurry was chosen because of its low water loss and good expansive
characteristics. This material cost less than one which did not use a
blend of fly ash.
A decision was made to begin the grouting operation through
grout hole No. 1. A check of the flow through the 2-inch pipe at this
hole showed a full stream of water emerging, but preparation continued
for using this hole to begin the grouting. The grout slurry was mixed
at a weight of 14.5 pounds per gallon using the impounded mine water.
The grouting started at a pressure of 10 psi. Approximately 400 gallons
of grout slurry were pumped through the pipe when a cement-colored dis-
charge from the 2-inch drain line was observed. The grout slurry was
flushed from the pipe with water and the valve was closed.
The water flow through the pipe in grout hole No. 2 was meas-
ured at a rate of 2.8 gallons per minute prior to beginning the remedial
grouting operation. The grouting was initiated at a pressure of 10 psi.
When 140 gallons of slurry had been pumped into the aggregate, the grout
broke through the main body of the coal, 10 feet in front of the aggre-
gate seal and 1-1/2 feet from the top of the coal seam. The grouting
operation was stopped at that point.
The flow of water through the pipe from grout hole No. 3 was
measured at 7 gallons per minute. The grouting again started at a
pressure of 10 psi. When 60 gallons of grout slurry had been pumped
into the pipe, water discolored with cement emerged through the mud
cavings 3 feet up on the coal seam at the front edge of the aggregate
seal. Heavy grout started flowing over the top of the aggregate about
one foot inside the right-hand row of posts. After pumping a total
of 260 gallons of grout, part of which overflowed the top of the aggre-
gate seal, the grouting was stopped.
The pipe in grout hole No. 4 did not have an initial water
discharge. Grouting began with a pressure of 15 psi. When 80 gallons
had been pumped into the pipe, grout started coming out the top part
of the aggregate seal. When 400 gallons of grout slurry had been
pumped in, the pumping was stopped.
The grouting operation then moved back to grout hole No. 1.
An additional 400 gallons of grout slurry were pumped into the pipe at
a pressure of 5 psi. There was a slight discoloration of water coming
from the mine drain line, but no other evidence of grout was observed.
-28-
-------
Another 400 gallons of grout slurry were pumped into the grout pipe at a
pressure of 10 psi. After this operation, the grouting was stopped.
A measurement of the flow from the grout pipes on the day after
the grouting operation indicated a flow of 0.35 gallon per minute from No.
1 grout pipe, 0.5 gallon per minute from No. 2 grout pipe and 0.1 gallon
per minute from No. 3 grout pipe, and no flow from No. 4 grout pipe. The
flow from the mine was 2.2 gallons per minute from the No. 1 Opening and
3.6 gallons per minute from the No. 2 Opening. This was a reduction of
about 17% in the total leakage from the mine.
Since a significant reduction in leakage was not obtained, a
decision was made to perform another grouting operation using a different
type grouting fluid. The grouting material selected was Halliburton's
DOC slurry. This slurry consists of cement particles surrounded by a
hydrocarbon base fluid incorporating a surface-active agent, or kerosene,
cement and a dispersant type surfactant. These materials are mixed to
form a slurry that remains inactive unless contacted by water. The
slurry will absorb water like a sponge to cause a hard and dense set.
This grouting operation was planned for grout holes No. 3 and 4
on Opening No. 2 because this opening had the larger flow coming from the
mine. Arrangements were made to divert the flow from the mine to recover
the kerosene and prevent discharge into the streams of the area.
The grouting was begun on grout hole No. 3. An initial flush
with kerosene was made at a pressure of 20 psi to remove any water from
the pipe to prevent contamination and premature setting of the grout.
The special slurry, weighing 15.4 Ibs/gal, used in the grouting operation
was mixed in 30-gallon batches and pumped through the grout pipe. Ten
batches were pumped through the pipe without observation of returns.
Grouting then began on grout hole No. 4 by pumping in 100
gallons of kerosene at 50 psi pressure as a spearhead, followed by a
30-gallon batch of the special grout. This was pumped into the pipe at
a rate of 20 gallons per minute. Cement returns were noticed immediately
coming over the top left of the aggregate seal, so the line was flushed
with kerosene. Another batch of the slurry was mixed at the same weight
and injected. The leak at the top of the aggregate was closed off, but
leakage was still observed on the lower right. Later, two other batches
were injected at a pressure of 150 psi without any cement returns being
observed. A batch was then pumped in at a weight of 15.6 Ibs/gal and
still no cement returns were observed. Another batch of 30 gallons was
pumped in with the pressure increased to 250 psi. No returns were seen.
An attempt was made to pump in another batch of 30 gallons, but the pipe
would not accept the grout slurry.
The connection was moved to grout hole No. 3 and the remaining
grout pumped in at this point, followed by another 30-gallon batch. No
visible returns were noted. Therefore, the grouting operation was halted
at this point.
-29-
-------
The flow rate on the two openings was measured on the follow-
ing day at 2.0 gpm for Opening No. 1 and 2.2 gpm for Opening No. 2.
This is a reduction of 27% from the flow prior to the second grouting
operation, or a 37% reduction from the original leakage.
Although the leakage had been reduced, it was decided to
try an additional grouting operation using the regular cement-fly ash
mixture which had been used in the first operation, with the addition
of Flocele, a shredded cellophane, as a plugging material. A total of
2,400 gallons of this material containing 75 pounds of Flocele was
pumped into Opening No. 1 through grout holes No. 1 and 2 at a maximum
pressure of 50 psi. A slight discoloration was noticed from the drain
line, but no cement grout returns were observed. The leakage from
Opening No. 1 after the grouting was measured at 1.6 gpm. On the follow-
ing day, the flow rate from Opening No. 1 was 2.12 gpm and from Opening
No. 2 was 2.41 gpm. There was essentially no reduction of leakage due
to this grouting.
Monitoring of the flow from the openings of Mine No. 40-016
was continued at regular intervals throughout the duration of the pro-
ject. The water data are shown in the Appendix (Table 5). It can be
noted that the level of the fluid impounded behind the seal is higher
in Opening No. 1 than in Opening No. 2, but the discharge rate is higher
from Opening No. 2. This could be a result of the bad condition of
Opening No. 2 prior to the original seal installation, particularly the
floor and roof areas, which would make a seal much more difficult to
obtain. The fluid level of impounded water increased over 10% during
the 11 months of monitoring after the remedial grouting. Two months
after grouting, the total leakage returned to a point approximately
equal to that measured prior to grouting, possibly due to the increased
head. However, before the grouting was performed, the average acidity
value for Opening No. 1 was 267 mg/1, while the average after grouting
was 150 mg/1. A similar reduction from 200 mg/1 to 96 mg/1 was noted
for Opening No. 2. This indicates that some channels of leakage around
the plug were sealed by the grouting and the leakage was channeled
through the limestone where some neutralization occurred.
The analyses of the samples from the two openings are listed
in the Appendix (Tables 6 and 7). All samples show a relatively high
acid and iron content, even when the pH varies from 3.1 to 6.7 from the
same opening. It should be noted that the pH increased after grouting.
The average pH of the discharge from Opening No. 1 increased from 4.1
to 5.3 and from 5.1 to 6.3 at Opening No. 2. This is in agreement with
the decrease in acidity mentioned above.
After grouting the seal and causing the water to flow through
the limestone, the acid production rate from Opening No. 1 decreased
from 64 pounds per week to 37 pounds per week. This is a reduction of
42%. In a similar manner, the acid production rate from Opening No. 2
was reduced 64%, or from 83 pounds per week to 30 pounds per week. The
-30-
-------
total acid production from both openings was reduced 54% by the grouting.
The cost for materials and equipment used in the remedial work
on Mine No. 40-016 was $6,007.00. The drilling and preparation of the
grout holes accounted for $1,468.00 of this cost. Site preparation and
restoration required an additional expenditure of $1,310.00. The cost
for the grout material pumped in place was determined. Exclusive of costs
for drilling and preparing grout holes and the site preparation and res-
toration, the cement grout cost $0.56 per gallon in place and the same
material containing Flocele was $0.57 per gallon. The DOC grout cost was
$3.41 per gallon in place.
-31-
-------
BULKHEAD DEVELOPMENT
Improvement of Aggregate Placement
During the previous study, research was conducted by the
Halliburton laboratories at Duncan, Oklahoma on movement of various
sizes of aggregate using air as the transporting medium. The aggregate
sizes used in the tests were 3/4-inch and 3/8-inch graded limestone.
The aggregate was moved from one pressure tank to another tank through
150 feet of 5-inch inside diameter pipe and hose. The pressure vessels
were 200-cubic-foot tanks with a rated working pressure of 40 psi. A
compressor rated at 465 scfm at 40 psi was used for the air supply.
The 3/4-inch limestone was conveyed from tank to tank at a
rate of approximately 600 pounds per minute at a pressure of 20 psi.
The 3/8-inch aggregate rate was 627 pounds per minute. This rate was
only for the actual conveying and did not include any time for reload-
ing. The rock was also conveyed a distance of 23 feet vertically dur-
ing these tests.
Additional tests were conducted at Weston, West Virginia using
truck mounted pressure tanks. These tests did not establish a rate, but
did confirm the feasibility of using this equipment for a field test.
A field test was performed using the two openings of Mine No.
40-016. For the placement of the aggregate in these openings, a 150-
cubic-foot truck mounted pressure tank was utilized. The tank was
loaded by a belt conveyor. The conveyor was supplied by a front-end
loader from a pile of aggregate on the work site. Loading of the tank
was accomplished at a rate of 1,000 pounds per minute for the 3/4- to
1-1/2-inch aggregate. An overall average of 500 pounds per minute was
pneumatically conveyed into the mine drift. An initial air pressure of
35 psi was used in the pressure tank. This pressure reduced to 10 psi
as the aggregate moved through the line.
In addition to the time required for loading which stopped the
conveying, a problem was encountered when trying to move wet aggregate.
The limestone movement through the pipe was decreased in proportion to
the fineness of the aggregate and the moisture present. It was found
that the minus 3/8-inch aggregate would tend to bridge because the mois-
ture would cause the fines and small rocks to bind together and prevent
movement, thereby forming a plug in the lines. It was also determined
that the rate and distance any aggregate could be moved was in inverse
proportion to its size.
The purpose of this phase of the project was to develop a
system for placement of the aggregate into a mine in a more efficient
manner than accomplished in the prior field tests. This was to be done
by developing a continuous handling system or a system with a minimum
downtime required for reloading.
-33-
-------
The first tests conducted were with 1-1/2-inch (No. 3) aggre-
gate. The pressurized tank was used without any modification. The
aggregate was loaded in one 160-cubic-foot pressure tank and attempts
were made to blow the material to an adjacent tank through 20 feet of
4-inch line. Air pressure of 30 psi was supplied by one compressor of
175 cfm capacity. Results were poor, so an additional compressor of
like capacity was added. The rock still did not move well. Additional
air was added up to 675 cfm. Rock moved slowly and bridged frequently
in the line, so testing of this size rock was discontinued.
A special bottom discharge manifold was then built to fit the
bottom of the pneumatic tank used in these tests. This manifold is
shown in Figure 15. It incorporated a connection for an air jet which
permitted air to be injected at the point where bridging was most likely
to occur.
New tests were arranged, using a 160-cubic-foot pressure tank
which had been modified by attaching the special bottom discharge manifold.
A view of the test site is shown in Figure 16. The aggregate was moved
back and forth between the two tank units. The amount moved was deter-
mined by weight of the unit on the left, which was located on a scale.
Thirty tests were made using conveying lines of 40, 60 and 80
feet, with initial tank pressures of 15 and 20 psi. Conveying rates
varied from 128 to 1,081 Ibs/min with an average rate of 604 Ibs/min.
This inconsistency in the rate, particularily at the conveying length of
80 feet, was due to several reasons. The tests were made over a period
of seven days under various weather conditions, including freezing
temperatures which caused the aggregate containing moisture to freeze
and bridge in the tank discharge. The first 23 tests were made using
the same batch of aggregate. Each test caused some wear on the rock
and produced more fines. The finer aggregate then moved at a faster
rate. Some of the rate variation was due to changes in compressor speed,
which altered the air output accordingly. Higher rates were obtained
from one unit then when conveying from the tank on the other unit, show-
ing that the tank condition affects the discharge rate. Detail test
data are shown in Table 8 in the Appendix.
An investigation was made to determine if the pneumatic con-
veying of the aggregate could be a continuous process over a long period
of time. A method using a hopper over an air line was considered, but
the idea was discarded when it was determined that previous tests had
already proven this method to be ineffective. A vane-type feeder was
also considered, but was rejected because of air leakage around vanes
and large aggregate size which created further sealing problems at the
vanes.
Since a method of continuous operation did not seem to be
feasible, plans were made for a test which would have the least possible
downtime. The test equipment included two mobile units having 160-cubic-
-34-
-------
CO
in
i
A HOT FORM LOC. 3 INTO £LLIPS£ AS SHOWN.
LENGTH OF ELLIPTICAL SECTION TO B£ B.OIM.
A TORCH CUT S.O Of*. HOLE IN LOC. 3 MO FIT + AS SHONN.
SMY VINYL/
HF-12.2
TQ.+38T5
'CZ2 I
52.5:353 I
f£U*LE HALF- +"XOOOm-IHECO INIH6 UNION
uuT-nex-it-a NC
HASHER-LOCK-t -STL. PL
FLANGS-SOCKET HVEUD-S
Z7.OOB99
3O.S3030
3/.SOOO+
13. IS OS
'-i°if°7
COUPLING -U NOM ll-j VLP
STL.
STL PL -ftSTM A36-1
STL PlPS-STOIHT-54 -IO.O
ELL INEL.D-+S"-S" STO WT-I.ON9KAO.
3
•i jii *
VE-OEMCO-S
MAN/FOLD-
MINK
FIGURE IS - BOTTOM DISCHARGE MANIFOLD - PRESSURE TANK
-------
I
CO
en
FIGURE 16 - PNEUMATIC CONVEYING TEST - 3/4-INCH AGGREGATE
-------
foot pressure tanks. The loaded tank was then moved forward so the
second unit could move under the conveyor to be loaded while the first
tank contents were being blown into the receiving trailer. The process
was then repeated, reloading the first tank while the second tank dis-
charged.
The test was initiated with the first tank containing 14,570
pounds of 3/4-inch aggregate and the second tank containing 11,200
pounds. Material was conveyed from the first tank through 80 feet of
4-inch hose to the receiving trailer in 25 minutes. Initial tank pres-
sure was 25 psi and 11 to 12 psi was maintained during time of dis-
charging with one 125-cfm compressor. The other two compressors were
discharging into the bottom discharge manifold.
Six minutes were required to make the necessary connections
to the second tank and initiate movement of aggregate. The first tank
was reloaded while unloading the second tank. Eight minutes were re-
quired to reload the first tank with the belt conveyor. Twenty-four
minutes were consumed in discharging aggregate from the second tank.
The reloaded tank was then discharged, requiring six minutes
for connections to be made and 15.65 minutes to move 11,150 pounds of
aggregate through the 80 feet of 4-inch hose.
In this test, 36,920 pounds of aggregate were moved 80 feet
pneumatically in 64.65 minutes with 12,minutes downtime between runs.
The total time of 76.65 minutes gives an equivalent rate of 481.7 pounds
per minute for the entire test.
One additional test was made using this procedure while build-
ing a bulkhead in a simulated mine fixture. This test is described in
another part of this report. During this test, six tank loadings were
made using a total of 66,220 pounds of 3/8-inch (No. 12) aggregate.
Conveying time for the six tanks of aggregate was 79.6 minutes. Time
for loading was only 24.3 minutes, or about 4 minutes average per tank.
The time for making connections was 8.7 minutes per load. This test
moved the aggregate at a rate of 544 pounds per minute. This is a
faster rate than was obtained in tests with the larger limestone.
Tests were conducted at tank pressures of 15 psi and 20 psi.
Results confirmed prior experience of Halliburton. As pressure in-
creased, the conveying rate of the aggregate increased. However, for
each size conveying line a maximum pressure was obtained for the max-
imum flow through that size line without stoppages. Increasing the
pressure results in bridging of the flow. The use of a larger convey-
ing line would permit higher tank pressures and a subsequent higher
discharge rate.
Moisture in the aggregate would reduce the discharge rate and
tend to clog the conveying line. Attempts to measure the moisture con-
-37-
-------
tent accurately were relatively unsuccessful since a representative
sample was difficult to obtain and transport to a laboratory without
losing moisture. A field sample sent to the Duncan laboratory showed
a moisture content of 3.57%, but the accuracy is questionable due to
the time and distance involved.
Tests conducted show that the conveying rate varies inversely
in proportion to aggregate particle size. Limestone was tested from
1-1/2-inch to 3/8-inch in size to obtain this information. The 3/4-inch
size aggregate was moved 80 feet during these tests and up to 150 feet
in previous tests. Smaller aggregate can be moved even further.
Since there are several variables in pneumatic conveying, the
movement of aggregate can be accomplished almost as desired by adjusting
the variables to meet the prescribed conditions.
Quick-Setting Mine Bulkheads
The purpose of this phase of the project was to develop some
type of bulkhead which could be placed remotely, used under variable mine
conditions, and offer a significant reduction in mine sealing costs.
The use of an expanding foam material was proposed as tempo-
rary front and rear bulkheads to contain a low-cost filling material.
A feasibility study was then conducted to investigate organic foam ma-
terials such as polyurethane and urea-formaldjehyde.
A polyurethane foam (Foam A) was considered, based on available
published literature and company experience on other projects. Chemical
and physical properties appeared to be satisfactory (Table 9 - Appendix).
However, uncontrollable mine atmosphere and temperature could adversely
affect the density of this foam. Simulated test equipment required was
studied and found to be expensive. Material costs were estimated to be
$1,000.00 fop'each of the two necessary bulkheads to contain a filler
material in constructing the seals, assuming the foam was applied at 3
pounds per cubic foot density. Polyurethane foam was not considered fur-
ther as a bulkhead material in mine sealing applications.
Urea-formaldehyde (Foam B) properties were considered and are
also listed in Table 9. Such foam material was developed primarily for
use as thermal insulation in building construction. It is less dense
and has lower compressive strength than polyurethane foam. Also, urea-
formaldehyde resin foams are susceptible to biological and dilute acid
attack, as well as being destroyed by contact with water. Foam generat-
ing units, available from several manufacturers, range in price from
$2,500.00 to $3,000.00. Although biological, acid, and/or water attack
would probably not be swift enough to prevent urea-formaldehyde foam
from serving as a temporary bulkhead, these factors together with low
strength, density and high equipment costs were responsible for abandon-
ment of further consideration of the foam in this application.
-38-
-------
Inorganic foam materials were investigated in the laboratory
and several formulations were tested (Table 10 - Appendix). Formal-
dehyde, hydrogen peroxide and sodium silicate were used in each case
together with a filler material such as silica flour or fly ash. The
first foams prepared were very fragile and crumbled easily under finger
pressure. Although foam strength increased AS the investigation pro-
ceeded, water permeability was found to be high. Also, hydrogen gas
was liberated in all formulations and could create potentially ex-
plosive mine atmospheric conditions. Formaldehyde could cause uncom-
fortable working conditions for personnel. No further consideration
was given to these inorganic foaming materials.
Research effort was then directed toward the development of a
material which would be quick-setting and self-supporting with sufficient
strength to withstand expected forces. Initial work was centered around
the use of various types of accelerators with cementitious materials.
Previous experience with cementing materials dictated the use
of bentonite in the mixtures to obtain better thixotropic and viscosity
characteristics. By prehydrating the bentonite and mixing it with a
slurry of cement, flocculation would occur and give higher viscosity.
Gypsum was known to achieve a fast set and impart viscosity, so it was
decided to incorporate gypsum into the tests.
Various slurries were prepared in the laboratory using com-
binations of portland cement, fly ash, gypsum, bentonite, calcium
chloride and salt. Ratios of water and other materials were adjusted
in an attempt to obtain the most desirable qualities. Twenty-four
different slurries were tested in the initial laboratory tests.
Slurries were checked for mixing characteristics, angle of repose and
set time. All of the slurry compositions tested and reasons for either
rejection or selection for further consideration are listed (Table 11 -
Appendix).
Eight slurries were selected with characteristics which would
be desirable for use in construction of a bulkhead or seal in a mine
drift. These slurries were given additional laboratory mixing tests
and compressive strength tests. Corapressive strength specifications
were set for a strength of 10 psi at 70°F within one hour after mixing,
and a strength of 300 psi at 70°F after 48 hours. This data is shown
in Table 12 in the Appendix.
Because of the quick set of the slurries, it was necessary to
depend on visual examination to determine how viscous each slurry was as
it was mixed. Slurry No. 13 appeared to be the most viscous, but was
expensive and would have contaminated water with diesel oil used in the
slurry and did not have the required compressive strength. Slurry No. 9
was viscous but did not have the required compressive strength. This
would have made the control of the pumping and mixing very difficult.
Slurries No. 12, 21 arid 22 were also considered sufficiently viscous.
-39-
-------
Cost for materials, exclusive of any charges for handling, blending,
mixing or pumping, ranged from $14.00 per cubic yard for Slurry No. 9
to $32.43 for Slurry No. 6. From these slurries, Slurry No. 22 was se-
lected for larger scale testing. This slurry was composed of 60 parts
cement, 40 parts gypsum, 6 parts bentonite and 100 parts water by weight,
and had a cost of $18.00 per cubic yard for materials.
In a simultaneous research effort, the use of portland cement
with various combinations of sodium silicate, silica flour, bentonite,
sand and fly ash was being studied (see Table 13 - Appendix). The pri-
mary problem encountered with these slurries was the high degree of
fluidity which caused a low angle of repose. It can be noted that for-
mula 1414-13-1 costs about one-third the cost of formula 1414-18-1. How-
ever, the most expensive formula develops sufficient strength to support
the weight of a man in less than one minute, while it requires several
hours for the other slurry to develop a similar strength.
Tests were initiated with full-size pumping and mixing equip-
ment using slurries developed in laboratory tests. The slurries were
designed as two-stream systems, so a special mixing head previously
designed by Halliburton engineers was used to blend the two slurries.
This head was a 5.5-inch diameter pipe, 18 inches long, with a 9-inch
diameter jacket around it. Slots in the inner pipe enabled fluid enter-
ing from each side of the jacket to mix as they entered. The manifolding
arrangement on the tests was such that the two fluids could be mixed to-
gether before they entered this jacket to provide longer mixing time.
This was the procedure used in the initial tests. Both the gypsum cement
slurry and the sodium silicate cement slurry were tested in this system.
The first tests indicated that sufficient mixing was not being
obtained through the mixing head. The mixing system was changed from
2-inch piping to 1-1/2-inch piping. The mixing head was removed. A
perforated disc was placed in the line so the slurries would pass through
it immediately after the two streams blended. Various lengths of pipe
from 5 feet to 63 feet were then added to the discharge system during the
tests. A better buildup of material was obtained with the longer pipe
which afforded additional mixing.
Twenty-eight full-scale test series were conducted in develop-
ing the correct slurry, placement equipment and techniques. Various ways
of mixing the components were tested for both the gypsum cement and
silicate cement slurries.
After twelve test series, an evaluation of the two general
types of slurries revealed the following information:
(1) The gypsum cement type slurry had excellent
viscosity properties and built up readily into
a wall. However, the slurry did not set fast
enough. When the wall reached a height of about
12 inches, the weight caused the lower part to
-40-
-------
slide. Figure 17 shows the good viscosity
characteristics of this slurry.
(2) The silicate cement slurry set very quickly,
but would not develop a viscosity sufficient
to allow a wall to be built.
In view of the above findings, it was decided to try to com-
bine the good viscosity qualities of the gypsum cement and the fast set
of the silicate cement. This resulted in many more tests, varying the
amount of sodium silicate, the percentage of bentonHe and the volumetric
ratio of the cement slurry to the bentonite-sodium silicate slurry. It
was also determined that the amount of mixing and the length of time
between mixing and discharge had a decided effect on the end product.
It was found that viscosity could be controlled either by changing the
length of pipe after the mixing or by varying the pumping rate. This
proved to be an excellent control factor when building a bulkhead.
The full-scale tests also were an aid in developing the proper
mixing and discharge equipment. After eight test series, the discharge
was changed from open-end tubing to a flattened nozzle about 1 inch by
4 inches, as shown in Figure 17. This nozzle plugged when using high
viscosity slurries. Two other variations of the flattened nozzle were
tested with similar results. Three additional nozzles were designed
and tested to provide a good discharge pattern and eliminate plugging
problems. Two of these were fabricated from tubing weld elbows. The
third nozzle was machined. A test of the machined nozzle, shown in
Figure 18, shows the excellent discharge pattern and coverage obtained.
This design was selected for use in the remainder of the laboratory and
field tests because it could be used in a 360° arc, it sprayed material
in a straight path as it was rotated and it showed no tendency to plug
with heavy slurries. A drawing of this nozzle is included as Figure 19.
After five additional test series had been made, a suitable
slurry had been developed composed of cement, bentonite, sodium silicate
and water that had an instantaneous viscosity sufficient to build up
properly and a set time of approximately 12 seconds. The ben ton ite and
sodium silicate slurry were mixed in a two-stream system with cement
slurry to produce the finished slurry. The cement slurry was cement and
water with a water-cement ratio of 0.43 or 4.85 gallons per sack. The
other slurry was composed of bentonite, sodium silicate and water. The
bentonite was 10% by weight of water used. It was prehydrated into the
water and then the sodium silicate was added. Tests proved that a
satisfactory ratio of bentonite slurry to sodium silicate was 3 to 1,
or the volume of sodium silicate would be one-third that of the ben-
tonite slurry. When the two slurry streams were mixed, the resultant
slurry characteristics varied according to the proportion of the mix-
ture. A greater amount of cement slurry caused higher viscosity while
an increase in the silicate gel slurry caused a faster set. After
evaluation of the test results, it was determined that a 1 to 1
-41-
-------
volumetric ratio-of the two slurries would produce the best effect.
A compressive strength test of the selected slurry indicated
a strength of 100 psi in 2 hours and 555 psi in 48 hours. The cost of
this slurry for materials only was calculated to be $20.90 per cubic
yard. Detailed data for this slurry are shown in Table 14 in the Appendix.
A simulated mine drift was built for a full-scale test. The
drift, constructed of wooden framing and plywood, was 10 feet wide, 7
feet high and 8 feet long. Three wooden posts were set inside the drift
on either side. A portion of the simulated mine drift is shown in
Figure 20. This picture also shows the beginning of the bulkhead which
was built in the opening.
The slurry used for this bulkhead was the selected sodium-
silicate type cement developed in previous tests. A batch of 15 cubic
yards was mixed, using the following components:
Slurry 1 Slurry 2
Water - 870 gallons Water - 1,125 gallons
Cement - 180 sacks Bentonite - 945 pounds
Sodium Silicate - 375 gallons
The two slurries were blended in equal volumes to produce the
finished slurry. Initial sample testing indicated that the viscosity
was too great in Slurry 2, so it was diluted with water. The final mix-
ture contained 8.2 per cent bentonite instead of 10 per cent. Future
work was done using 8 per cent.
The two slurries were pumped with Moyno pumps. An equal
volume was maintained from each pump. This was accomplished by prior
pump tests to calibrate the pumps so the volume could be known by read-
ing the pump speed. The volume pumped was indicated on special meters
having dial faces reading in gallons per minute.
The length of pipe after the two slurries blended was 36 feet.
The placement of the slurry was controlled by an operator moving the
pipe to turn the nozzle from side to side. Figure 20 illustrates how
the slurry was placed by oscillation of the pipe. Figure 21 shows the
completed bulkhead which completely filled and sealed the simulated drift.
Following the successful completion of the bulkhead in the
simulated mine drift, preparation was started to build a similar bulk-
head in an abandoned mine drift. Mine No. 62-008 was selected for this
field test.
Before any field tests could be made, it was necessary to secure
access rights to the properties involved. Obtaining these access rights
usually requires considerable time and money. Legal services are normally
required. In some cases, a large number of owners are involved, each
-42-
-------
u!
FIGURE 17 - BUILDUP OF GYPSUM CEMENT SLURRY IN YARD TEST
FIGURE 18 - NOZZLE DEVELOPMENT TEST
-43-
-------
secr/ow A A
z s+.tosos i srt gAx-cr/t/s/ toie • tt /to -e.f
FIGURE 19 - SPECIAL NOZZLE FOR HIGH VISCOSITY SLURRIES
-------
FIGURE ZO-SLURRY DISTRIBUTION IN BULKHEAD CONSTRUCTION-SIMULATED MINE TEST
-45-
-------
FIGURE 21 - COMPLETED BULKHEAD - SIMULATED MINE TEST
-------
of whom must be located to complete the access agreement.
Mine No. 62-008 is a small abandoned drift mine complex located
3 miles west of Lost Creek, Harrison County, West Virginia. Twenty acres
of Pittsburgh coal have been deep-mined and approximately 130 acres re-
main in reserve. The Redstone coal contains a small drift mine and has
been augered along with the Pittsburgh seam. The periphery has been
stripped for both the Pittsburgh and Redstone coal, leaving a steep con-
tinuous highwall.
A mine map of this mine is attached as Figure 22. Three open-
ings are evident in the Redstone coal, but only seepage was observed from
thes.e openings. Four openings were observed in the Pittsburgh coal and
three were draining slightly. The fourth opening was a 4-foot diameter
airway which was partially collapsed and dry. The flow from these three
openings was measured previously at 5 gpm. Sample analysis indicated a
pH of 2.85 and an acidity of 459 mg/1, or 27.5 pounds per day of acid
production. Analysis also indicated an iron content of 119 mg/1, or 7.1
pounds per day of iron, and 3500 mg/1 of sulfate. This water is being
discharged into a tributary of Lost Creek. The coal seams dip steeply
to the east. Three fresh water impoundments were found on the site. A
geologic section of this mine is attached (Figure 65 - Appendix).
The opening selected for the first field test was the main por-
tal, denoted as Opening No. 4. A view of this opening is shown in Figure
23 before starting any cleanup or preparation. This opening was partially
open and had good roof conditions. Figure 24 shows the openings in Mine
No. 62-008 after site was cleaned, openings enlarged and excessive backfill
removed. The main portal is the left opening in this picture. Drift is
5 feet high and 11 feet wide.
Figure 25 shows the inside of the main portal before cleanup.
The portal included posts in the interior similar to those in the sim-
ulated mine test. After cleanup, the floor, walls and roof were jetted
with water to remove mud and slime to provide a good bonding surface for
the cement bulkhead. The jetting was accomplished by pumping with a
high-pressure pump through the pipe and nozzle on the two-wheel dolly,
which was later used to place the slurry in the mine for the bulkhead.
A 2-inch plastic line was placed on the floor to extend from behind the
bulkhead to serve as a drain line.
The two slurries were mixed in 250-cubic foot cone-bottom
tanks. The bentonite was prehydrated by pumping water through the reg-
ular Halliburton mixing system and slowly adding the bentonite. After
this was mixed and stirred by air for 30 minutes, sodium silicate was
added through the same system and mixed with air in the bentonite slurry.
The cement was mixed by blowing bulk cement from a pressurized tank into
the tank of water and stirring with air jets. The temperature of the
cement slurry was 67°F and the weight was 15.4 Ibs/gal. The temperature
of the gelled silicate slurry was 45°F.
These slurries were again batch-mixed, using the following
proportions:
-47-
-------
62-008 (Pittsburgh Coal)
FIGURE 22 - MINE MAP - MINE NO. 62-008
-------
FIGURE 23 - MAIN PORTAL - MINE NO. 62-008 BEFORE CLEANUP
FIGURE 24 - WORK SITE AFTER CLEANUP - MINE NO. 62-008
-49-
-------
01
c::>
. ..in.*-
FIGURE 25 - INSIDE OF MAIN PORTAL BEFORE CLEANUP - MINE NO. 62-008
-------
Slurry 1 Slurry 2 ^__
Water - 816 gallons Water- 1,050 gallons
Cement - 168 sacks Bentonite - 700 pounds
Sodium Silicate - 350 gallons
Sample tests mixing the two slurries showed that the resultant
slurry had sufficient viscosity, but would not set quickly. It was
thought that the cause for this was the low temperature of the slurries,
particularly the bentonite-silicate slurry. To offset this, additional
sodium silicate was added for the sample tests. This proved successful,
so an additional 165 gallons of sodium silicate was added to Slurry 2.
These slurries were then used to blend together for the bulkhead con-
struction.
The building of the bulkhead was started by pumping equal
volumes of each slurry, bringing them together in a manifold just out-
side the mine. The blended slurry was then pumped through a perforated
plate in the discharge pipe and on into the drift. The beginning of the
bulkhead construction is shown in Figure 26. The completed bulkhead in
this drift is shown in Figure 27. Fifty minutes were required to build
the bulkhead after starting to pump the two slurries into the mine drift.
The angle of repose on the front side of the bulkhead was about 60 degrees
from the horizontal. A drawing of a section through this bulkhead is
attached as Figure 28.
It was discovered during the two full-scale tests that there
were several critical points. The temperatures of the slurries for the
mine bulkhead were low, causing a problem in getting the proper setting
time. The technique of proper placement td prevent sloughing was found
to be very important, as well as maintaining the proper pump rates and
slurry ratio.
Because of the low temperature problem, tests were conducted
in the laboratory using materials at 40°F. Slurries were mixed using
bentonite-sodium silicate ratios of 2 to 1 and 3 to 1. Bentonite
ranged from 8 per cent to 12 per cent. Some tests included small
amounts of sodium bicarbonate. It was found that when the temperature
was low, especially below 50°F, it had an influence on the setting
properties of the final s,lurry, regardless of the amount of bentonite
used or the degree of its hydration. The effect was more pronounced
as the per cent of bentonite was reduced. The data indicates that a
bentonite-sodium silicate ratio of 2 to 1 rather than 3 to 1 was
necessary when the temperature was 40°F or 50°F (Table 15 - Appendix).
Full-scale tests were again conducted in the yard by spray-
ing slurry on the ground for observation. The effect of using a new
mixing tube to replace the perforated disc was evaluated and found to
provide excellent mixing with considerable reduction in pressure re-
quirement. This mixer nipple is an 18-inch long section of 1-1/2-inch
pipe with alternate horizontal and vertical thin section baffles. These
act as shear plates to impart shear to the fluid which acts to mix or
blend all parts of the slurry. A drawing of this mixer nipple is shown
in Figure 29. This type mixer nipple was used in the discharge line
-51-
-------
FIGURE 26 - BEGINNING OF BULKHEAD CONSTRUCTION IN MINE NO. 62-008
FIGURE 27 - COMPLETED BULKHEAD IN MINE NO. 62-008
-------
5'10"
AV6.
. \ . •»• • » .
. ' • • « « .
» »
-10'.
AVG.
SECTION-QUICK SETTING BULKHEAD-OPENING N0.4-MINE NO. 62-008
FIGURE 28
-------
Ul
A
J
ttl
— XO KtK
DETAIL. LOC.
STAMP PART NO. -
SECTION A\ A\
SJ.gJg Z \STL. fH-HFO.KHtftZ
-------
on all further testing.
The full-scale tests also confirmed the laboratory tests in
regard to temperature. Slurries tested at low temperatures did not
set quickly. The slurry would spread out before a set would occur.
These tests also confirmed that the temperature should be at least
65°F in order to enable the slurry to set quickly and build up prop-
erly. Tests were also made by mixing the slurries with heated water.
Water was heated to 78°F prior to mixing with cement or bentonite
Using the original ratios of bentonite slurry to sodium silicate of
3 to 1, a slurry was produced which set quickly and built up into a
wall with a high angle of repose. Figure 30 shows a view of this
test. Two mixing nipples, providing 3 feet of length, were used on
these and subsequent tests. The length of pipe used in the discharge
line after the two slurries blended did not seem to be very critical.
Lengths of 8 to 20 feet were used in most tests, but variations up to
30 feet did not change results appreciably. The controlling factor
used in the tests which seemed most effective was the variation in
pump rate and in the ratio of the two slurries. These were used as
control methods when building bulkheads.
The yard tests indicated that there was little advantage in
using 10 per cent bentonite if the water temperature was 65°F or higher,
so 8 per cent was used as standard for further mixing.
Laboratory tests were made on slurry composed of 50%
cement and 50% fly ash with calcium chloride as the accelerator.
This material would not be less expensive than the silicate cement,
but the mixing procedures would be simplified and therefore more
economical. The tests showed that the slurry might be able to sat-
isfactorily build a bulkhead, since the set time could be designed
as low as 1-1/2 minutes with another minute required for self-support.
However, full-scale yard tests proved that neither the set time nor
the self-support time was short enough to permit the material to build
up into a bulkhead, so this approach was abandoned.
Further field testing of the bulkhead construction materials
and technique was scheduled for a drift opening in Mine No. 62-008 near
Clarksburg, West Virginia. This portal, adjacent to the main opening,
was designated as Opening No. 5. It was planned to build a seal in
this opening by placing a front and rear bulkhead of quick-setting
material and filling between the bulkheads with an expansive grout
slurry. Flow of water from the drift prior to this test was 15 gpm.
An analysis indicated that the discharge had a pH of 2.8, acidity
of 2260 mg/1 and iron content of 600 mg/1. This amounts to a daily
production of 413 pounds of acid and 100 pounds of iron.
The equipment used for placing this seal in the mine was the
same as used previously to build the bulkhead in an adjoining opening
(see Figure 24), with the addition of a large tank which was used for
-55-
-------
en
-t-
FIGURE 30 - YARD TESTS OF SLURRY USING HEATED WATER
-------
heating the water. The work was conducted when the temperature was about
21°F and snowing. Water was heated to 62°F to mix the slurries. Cement
used was API Class A. Eight per cent bentonite was used with a ratio of
3 to 1 for the prehydrated bentonite-sodium silicate slurry. A volume of
12 cubic yards was prepared for the rear bulkhead.
The dolly which was used to carry the discharge pipe and slurry
lines permitted the operator to work remotely from the place where bulk-
head was built. Figure 31 shows this equipment in use on the mine seal
construction. There were 17 feet of discharge pipe ahead of the swivel
where the two slurries blended, including the two mixer nipples. The
dolly contained two electric lights mounted just behind the nozzle for
illuminating the discharge area.
When the operation of placing the slurry began, the heat from
the slurry caused a fog around the bulkhead. This made it difficult for
the operator to see the progress of work, but did not cause any insurmount-
able problem. The construction of the bulkhead was complete in less than
one hour and required 12 cubic yards of slurry.
After this bulkhead was completed, 2-inch plastic pipes were
installed at floor and roof from the area in front of the bulkhead to the
outside entrance and fitted with valves to control the grouting operation.
These were used for grouting the interior part after constructing the
front bulkhead. Figure 32 is a schematic drawing showing grout pipe
locations.
The front bulkhead was built in the same manner as the rear bulk-
head. Water temperatures were 59°F for the bentonite slurry and 62°F for
the cement slurry. The ambient temperature was 20°F. Since the temper-
atures of the slurries were slightly below the desired minimum, special
care was given to material placement, slurry ratios and pump rates to
assure the material setting sufficiently to support its weight. The con-
struction of this bulkhead was completed in slightly over one hour after
materials were mixed and required 10 cubic yards of slurry.
On the following day, water containing a dye was pumped into
the void between the two bulkheads to determine its volume and spot any
leaks. When 4,410 gallons had been pumped into the void, a leak was
observed around the upper left grout pipe. With a volume of 5,040
gallons, another leak occurred around the upper center grout pipe.
Pumping was discontinued. The volume of the void was determined to be
773 cubic feet, based on volume of water required to fill to level of
grout pipes and calculated volume above that point.
The void was filled by pumping Halliburton LIGHT Cement through
the bottom grout pipe. This grouting material was a blend of 65% port-
land cement, 35% fly ash and 6% bentonite with 9.9 gallons of water per
bulk cubic foot of dry material, or 61 parts cement, 26 parts fly ash,
5.2 parts bentonite and 82.5 parts water by weight.
-57-
-------
When 478 cubic feet were Injected, the grout flowed from the
upper left grout pipe. Valve was closed and pumping continued until
773 cubic feet were injected. At this time, leakage was observed across
a large portion of the top of the bulkhead. The cement slurry was drained
until the level was below the upper grout pipes and then allowed to harden.
The second stage of grouting was accomplished by pumping the
grout through the upper right grout pipe. The same grout slurry was
used as in the first stage with the addition of sawdust, shredded cello-
phane and shredded cane fiber. This grout containing the additives
successfully sealed the leaks. A 4-foot head was placed on the slurry
while it hardened by a riser pipe attached to the front of the center
upper grout pipe. The drain line was fitted with a manometer to meas-
ure the head of water which would form behind the mine seal. A draw-
ing of a section through this mine seal is attached as Figure 33.
The three bulkheads built of the sodium silicate, bentonite
and cement mixture have proven to be excellent seals. A schematic
drawing showing these seals in place in Mine No. 62-008 is attached as
Figure 34.
Monitoring of the water head retained by the seals was begun
immediately after construction and continued throughout the project.
After 10 months, the single bulkhead in Opening No. 4 showed only slight
seepage and contained a head of 54.5 inches behind it. The double bulk-
head seal in Opening No. 5 after 7 months showed no leakage and contained
a head of 51 inches of water (Tables 16 and 17). There were no water
samples collected from Openings 4 and 5 after installation of seals be-
cause the drainage stopped.
The cost for the single bulkhead in Opening No. 4 was $3,564.00.
This included $647.00 for site preparation, $1,165.00 for materials and
$1,752.00 for equipment and operators. The construction cost for the
double bulkhead seal with filler material was $9,449.00. This included
$894.00 for site preparation, $3,872.00 for materials and $4,683.00 for
equipment and operators. This cost could probably be reduced when this
type of construction is performed on an operational basis.
-58-
-------
-:--
VD
''
•*•"
m • \^-*» ~~ -1
FIGURE 31 - REMOTE OPERATION OF EQUIPMENT FOR PLACEMENT OF SLURRY
• 3
A(*
/*
"; *
-------
O
6"-*-
V^-
k-
/
J— *
ALL PIPES ARE 2 INCH PLASTIC.
REAR
BULKHEAD
* I - LENGTH 19 FEET
*2*4- LENGTH 12.5 FEET
* 3-LENGTH 20 FEET
* 5-DRAIN PIPE-LENGTH 40 FEET
GROUT PIPE LOCATION-OPENING NO. 5 - MINE NO. 62-008
FIGURE 32
-------
t
5' 2"
AV6.
I
O • , ° ' < °
* * * t *
|
•
, •
* » t
. • • o 'o •
««»" ^
, D . •/ REAR• '
BULKHEAD •
22'6"-
12'10"-
35'4"-
SECTION-MINE SEAL-OPENING NO. 5-MINE NO. 62-008
FIGURE 33
-------
I
en
ro
*I2 LIMESTONE
PLUG
AUGER
MINED
SECTION
SINGLE
CEMENT
BULKHEAD
2 BULKHEADS WITH LIGHT WEIGHT
CEMENT BETWEEN
PARTIALLY COLLAPSED FAN WAY
SP-8
PORTAL
SP-4
UTILITY
PORTAL
SP-5
[-7 J I
HIGHWALL
OPENING SP-3
SCALE' I "-50"
FIGURE 34 - REMEDIAL WORK - MINE NO. 62-008
-------
PERMEABLE PLUG INVESTIGATION
Research was conducted to determine the feasibility of using
permeable plugs to provide for drainage of some mine water with in-place
treatment of the fluid as it passes through the plug.
Several approaches were investigated. One possibility was to
use some type of mine seal containing a conduction pipe. The pipe was
to contain either an electrode to remove the iron or a bed of ion exchange
resin which could be recharged to remove sulfate and iron. Another pro-
posal was made to use calcium phosphate, barium carbonate or phosphate
rock in a drainage pipe to treat the water. In this method, as materials
were expended they could be replaced with fresh chemicals. These methods
were rejected from further consideration because of the maintenance
problem involved.
The idea which seemed to have the most merit was that of
pneumatically placing a plug of graded aggregate in a mine drift. The
aggregate would be so graded that acid mine water would flow through the
plug with sufficient retention time to be neutralized. As iron hydrox-
ide (and possibly calcium sulfate) resulted from the neutralization
process, the pores would gradually fill so that the result would be a
plug which would seal the mine drift.
Two different approaches were taken. The first idea was to
use an aggregate which would not seal quickly but would neutralize the
water flowing through the plug. The other approach was to use an
aggregate which would quickly form a seal.
Various grades of aggregate were investigated to find the
proper size distribution which might be adaptable to one of the desired
approaches. Screen analyses were made of samples from two vendors in
the project area. These vendors were Paul Harrold Limestone Company,
Clarksburg, West Virginia and Greer Limestone Company of Greer, West
Virginia. The screen analyses are attached in Table 18 of the
Appendix.
The permeability of a plug composed of graded limestone sub-
mitted by each company was determined. Permeability is a measurement
of the resistance to fluid flow offered by a known volume of particles,
usually expressed in darcies. Darcy's law, which states that the ve-
locity of flow of a liquid through a porous medium due to difference
in pressure is proportional to the pressure gradient in the direction
of flow, permits the determination of permeability mathematically.
Flow rate was determined while maintaining a constant pressure drop
under controlled laboratory test conditions through a definite size
sample of graded limestone using water as the fluid. These values
were then used in a computer program to calculate the permeability
using the following formula based on Darcy's law:
-63-
-------
K - 11 L Q (I)
A AP
where K = Permeability, Darcies
y = Viscosity, centipoise
Q = Flow Rate, ml/sec
L = Length over which pressure drop
measured, cm
AP = Pressure Drop, atmosphere
A = Cross-sectional area, sq cm
The permeability of the various limestone samples was deter-
mined to be as follows:
Permeability
Aggregate (Darcies)
Harrold #13 0.031
Greer #13 0.648
Harrold #12 1685.
Greer #12 4937.
Greer #12B 0.748
Harrold #10 8430.
Harrold #7 9535.
After a study of the screen analyses and permeability values,
Harrold #12 limestone was selected for laboratory tests for a plug which
would not seal quickly. Greer #12 and Harrold #10 and #7 had permea-
bility so high it was felt that retention time involved would necessitate
a plug too large to be practical.
The Greer #13 was modified by removing all particles that
passed a 40 mesh (U.S. Standard Sieve) screen. This removed about 25%
of the material in the sample. The modified Greer #13 material was then
found to have a permeability of 210 Darcies. This material was then
used for laboratory tests of the material to plug quickly. This mate-
rial was also tested with 1% barium carbonate added to evaluate the
effect of reduced permeability and to attempt to facilitate the pre-
cipitation of barium sulfate.
Formula I was used to calculate the length required for a lime-
stone plug having a cross section of 5 x 10 feet, when the pressure drop
was 2 psi and the flow rate 30 gpm. Assuming that the flow would only
go through one foot of depth initially, it was found that a plug with a
minimum length of 20 feet would be satisfactory for the Harrold #12
limestone.
-64-
-------
The retention time for the acid mine water in the limestone
bed was calculated using the following formula:
Retention Time, minutes =
where
T
V = Volume, cubic feet
$ = Porosity, per cent
(II)
Q = Flow Rate, cfm
The minimum length of 20 feet was assumed for the 5 x 10 foot
limestone plug. The porosity of the sample of Harrold #12 limestone
was determined in the laboratory to be 40%. The flow was established
at 30 gpm. The calculations showed the retention time for these con-
ditions to be 23 minutes.
limestone
A laboratory model was constructed using a 6-inch column of
i in a 1-inch glass tube. This apparatus is shown in Figure 35,
FIGURE 35
LABORATORY APPARATUS FOR PERMEABILITY STUDIES
-65-
-------
Flow rate of acid mine water through the Harrold #12 graded
limestone was set initially at 141 cc/hr to correspond to a retention
time of 23 minutes. A constant flow rate was very difficult to main-
tain in the laboratory test due to precipitation of iron hydroxide as
the water was neutralized; however, tests were able to confirm the re-
tention time as sufficient to neutralize the acid mine water. The var-
iation in pH is due to fluctuation in the flow rate. After 209 hours,
it was not possible to continue any flow through the test column (Table 19 •
Appendix).
Additional laboratory tests were conducted using the same test
apparatus to determine the length of time it would require to plug a lime-
stone column with less permeability. A constant head of acid water was
maintained to exert a pressure of 2 psi on the limestone column. Water
samples were analyzed at various times during the test. Flow rate was
set initially to correspond to a retention time of 23 minutes. This flow
rate gradually decreased during the test until no flow could be obtained
through the limestone due to precipitation of iron hydroxide.
The modified Greer #13 was tested. It completely plugged in
125 hours. The same material with 1% barium carbonate added was also
tested with similar results (Tables 20 and 21 - Appendix).
An analysis of the acid mine water used in these tests revealed
the following characteristics:
Acidity Alkalinity Total Dissolved
as CaCOs as CaCOs Iron Iron Ca++ 804
PH (mg/1) (mg/1) (mg/1) (mg/1) (mg/1) (mg/1)
3.2 180 0 30 25 500 1450
It can also be noted that in each test the acid mine water was
neutralized. The addition of the 1% barium carbonate to the modified
Greer #13 limestone reduced the permeability and caused an immediate de-
crease in flow rate as compared to a gradual decrease throughout the test
for the modified Greer #13. This caused slightly higher pH values with the
limestone-barium carbonate mixture, but did not increase the sulfate.
After evaluating the laboratory test results, it was decided
that a plug of the graded limestone should be placed in a drift opening
for field evaluation. A field test was made under controlled conditions
to see if data would confirm the laboratory results.
Mine No. 40-085 was selected for the site of the installation.
This was an open drift mine in Mount Clare, Harrison County, West Vir-
ginia which had not been affected by stripping. Water was flowing from
an 8 x 5 foot opening in excess of 20 gpm into Browns Creek. The pH of
the water was 3.3.
The ground was excavated to receive a wooden trough in the dis-
-66-
-------
charge flow path. The trough was constructed in the excavated area
using wood framing and marine plywood. The finished trough was 5 feet
high, 6 feet wide and 30 feet long. Three separate compartments 2 feet
wide were provided across the width, with 3-inch standpipes in each
compartment at the center and input end. Standpipes had the lower
30 inches perforated, enabling the height of water to be measured at
each point. The standpipe next to the inlet reservoir was designated
as number 1 and the pipe in the center as number 2. A prefix was added
to designate the compartment number. Thus, location 1-2 would be the
center standpipe of compartment number 1. The completed trough is shown
in Figure 36 before filling with limestone.
The drain end of the trough contained a 3-inch drain hole at
the lower part of each compartment. The trough was filled with Harrold
#12 limestone. A flume 16 inches square was made to bring the water from
the mine discharge into an input reservoir on one end of the trough. The
inlet connections were installed with a valve on each compartment to reg-
ulate the input flow. In Compartment No. 1, a 3-foot extension pipe was
placed below the valve so fluid discharged 2 feet above the bottom of the
trough in the limestone bed. In Compartment No. 2 (center) the fluid
was injected at the top of the compartment. In Compartment No. 3, a
2-inch plastic pipe 20 feet in length was placed on the top of the lime-
stone and connected to the inlet valve. Holes 1/2-inch in diameter were
drilled at intervals along the pipe in order to distribute 10 gpm of
acid mine water along the length of the limestone bed. Figure 37 shows
the inlet of the trough with the valves and inlet standpipes clearly
visible. A 4-inch line was installed above the top of the bed level as
an overflow line, and an 8-inch line installed to bypass surface water
around the installation.
Flow was initiated through the trough. It was quickly evident
that the water would not readily flow out the discharge end of the com-
partments, so the end of the trough was removed and the limestone allowed
to take a natural angle of repose. The top was covered with plastic to
prevent entry of water other than mine discharge.
The flow was then monitored through each compartment and at
each standpipe. The input flow was set at 9.4 gpm for each compartment
(Table 22 - Appendix). Figure 38 is a graph of the flow through each
compartment. It can be noted that the flow decreased steadily in each
compartment. Compartment No. 2 overflowed into adjoining compartments
on the twelfth day, and Compartment No. 1 overflowed the next day. At
this time, flow input on No. 1 was changed to top input similar to that
in No. 2 and input flow readjusted into these two compartments to pro-
vide maximum flow possible without overflow.
Throughout the balance of this test, it was necessary to clean
and adjust the valves periodically to maintain flow through the trough.
The water level in the limestone as measured by the stand-
pipes is also shown in Table 22. A comparison of the level in each
-67-
-------
compartment at several times during the test is graphically illustrated
in Figure 39. The numbers of the lines show the sequence of measure-
ments taken throughout the test.
It will be noted that in Compartments No. 1 and No. 2, the
flow pattern was very similar. The height of flow increased at stand-
pipe 1 and decreased at standpipe 2. In Compartment No. 3, the flow
was more uniform along the length of the trough than in the other two
compartments as indicated by the height decreasing in both standpipes.
More flow was required to effect a plugging in the limestone. Whereas
Compartments No. 1 and 2 plugged with precipitate in 12 and 13 days as
indicated by overflowing, Compartment No. 3 did not completely stop flow-
ing during the test. The flow through Compartments No. 1 and No. 2 after
plugging occurred was undoubtedly due to leakage from Compartment No. 3.
The test was stopped after 49 days. Evaluation of the test
was difficult since considerable leakage was noted from the wooden
trough and possibly from one compartment to another. In addition,
plugging was continually encountered in the three inlet valves as well
as the discharge holes in the plastic pipe used along the length of
Compartment No. 3. Since it was necessary to clean and readjust the
valves several times, it was not possible to maintain a constant input
rate. However, the graded limestone plug did gradually seal due to
precipitation of iron hydroxide which plugged the pore space in the
limestone.
In the compartments where the water was introduced into the
front end of the #12 limestone, the plugging occurred for the full
height in the first few feet adjacent to the water input point, and
gradually declined to approximately 6 inches height just before dis-
charge from the limestone. In the compartment where the water was
introduced along its length through a perforated pipe, the plugging
occurred for the full height throughout the 20-foot section covered
by the pipe. This can be noted on Figure 40. This drawing depicts
the plugged area of each compartment as determined by the stain from
the iron floe precipitated during the neutralization process.
It was also found that the water passing through the plug
was partially neutralized and iron content reduced. The average
acidity for the mine water in this test was 568.5 mg/1. The average
reduction in acidity by passage through each compartment ranged from
8}% for Compartment No. 1 to 89% for Compartment No. 3. The average
iron content for the mine water was 125.9 mg/1. The average reduction
in iron during the test varied from 48.5% for Compartment No. 1 to
53% for Compartment No. 2 and 3. The mine water pH of 3.2 to 3.8 was
changed to a pH of 5.7 to 7.0 by passage through the 30 feet of lime-
stone. Analyses of all water samples obtained in this test are in-
cluded as Table 23 in the Appendix.
A sample of Harrolds #12 limestone which was coated with sludge
was taken from the trough for analysis to determine the volume and com-
-68-
-------
FIGURE 36 - COMPLETED TROUGH - PERMEABLE PLUG TESTS
FIGURE 37 - TROUGH - INLET COMPARTMENT AND VALVES
-69-
-------
CLEANED
VALVES
. COMPARTMENT *l
0 COMPARTMENT *2
El COMPARTMENT *3
0 3 6 9 12 15 18 21 24 27 30 33 36 39 42 45 48
ELAPSED TIME,DAYS
FLOW THROUGH LIMESTONE IN WOODEN TROUGH-PERMEABLE PLUG TEST NO. I
FIGURE 38
-------
NOTEs LINE NUMBERS SHOW TIME SEQUENCE OF MEASUREMENT.
COMPARTMENT*! COMPARTMENT * 2 COMPARTMENT * 3
1-2
STANDPIPE NO.
2-1 2-2
STANDPIPE NO.
3-1
3-2
STANDPIPE NO.
WATER LEVEL IN LIMESTONE FILLED TROUGH - PERMEABLE PLUG TEST NO. I
FIGURE 39
-------
rva
COMPARTMENT NO. I - PERMEABLE PLUG TEST NO. I
COMPARTMENT NO. 2 - PERMEABLE PLUG TEST NO. I
INLET
INLET
COMPARTMENT NO. 3 - PERMEABLE PLUG TEST NO. I
FIGURE 40 - PLUGGING PROFILE - WOODEN TROUGH - TEST NO. I
-------
position of the sludge formed during the test. It was determined in lab-
oratory examination that 12 cc of sludge were deposited for each 100 cc
of limestone in the particular sample submitted, but this could vary some-
what depending on where and how carefully the sample was taken. Based on
this analysis, approximately 30 cubic feet of sludge would be obtained for
each compartment in 49 days. The X-ray analysis showed the sludge coating
to be primarily amorphous iron oxide with a small amount of quartz and
iron carbonate.
After the evaluation of these test results, it was decided to
conduct an additional test in the wooden trough. The purpose was to con-
firm the previous test results with #12 limestone and to evaluate a lime-
stone in the same manner which had a higher permeability.
In order to eliminate the leakage problem in the wooden trough,
the compartments were cleaned out and then sealed. Sealing was accom-
plished by grouting beneath the trough with cement slurry to fill all
voids, caulking all seams in the compartment with a caulking compound
and placing a 3-inch concrete floor in the trough. In addition, the com-
partments were lined with a layer of 6 mil plastic sheeting before fill-
ing with limestone.
Each compartment contained three sample tubes located strate-
gically along the length of the trough. These tubes were 3-1/2-inch OD
plastic 24 inches long, perforated with 1/2-inch holes at 6-inch centers
on two sides. Tubes were filled with limestone identical to that in the
compartment and ends capped. They were then placed in the compartment
in an upright position as the compartment was filled with limestone.
Gauging pipes were made of 2-3/8-inch OD plastic with 7/16-inch holes
drilled at 3-inch centers on two sides over the lower 30 inches. These
were placed at five points in each compartment with 5 feet between pipes.
A drawing showing a section of each type of compartment is attached as
Figure 41. Compartments No. 1 and No. 3 each had a roof over the lime-
stone at a height of 36 inches above the floor. Compartment No. 1 con-
tained Harrold #12 limestone and Compartment No. 3 held Harrold #7 lime-
stone. Compartment No. 2 contained Harrold #7 limestone to a depth of
about 4 feet. Two 2-3/8-inch OD plastic distribution pipes, 20 feet
long with 1/4-inch holes on 10-inch centers, were placed in the lime-
stone. One was 21 inches above the floor and the other was 47 inches.
Figure 42 shows a view of the wooden trough as used in this
test. The roof and protruding gauging pipes can clearly be seen in the
two outer compartments, as well as the limestone and gauging pipes in
the center compartment. Figure 43 shows the discharge end of the trough.
The large by-pass pipe is shown on the left side. The input arrange-
ment for the trough is shown in Figure 44. Note that the flow comes
from the input reservoir into a sump in each compartment, and then flows
through the limestone or into the distribution pipes. The amount of
flow desired can be obtained by adjustment of valves in the input reser-
voir. Any excess flow above that desired will go through the large by-
pass pipe on the outside of the trough.
-73-
-------
-S'O"
-S'O'-
-S'O"-
-sV
S'O"
.'OVERFLOW PIPE
NO. 7 UMESTONE_.
/•SAMPLE
r/ TUBE
S'O' —
-S'O'
^OAUOINO /
' TUBE /
DISTRIBUTION/
PIPES \
47"
l!
CONCRETE
COMPARTMENT NO. 2 - TEST NO. 2
OVERFLOW
EFFLUENT C±3j
-6'6"
'
ft "
- 5'0" -t» S'O"
} LJMESTONE;
COMPARTMENTS NO. I 9 3 - TEST NO. 2
SECTION-COMPARTMENTS-WOODEN TEST TROUGH
FIGURE 41
-------
' !
U i
VI. I
FIGURE 42 - TROUGH - OVERALL VIEW - TEST MO. 2
-------
FIGURE 43 - TROUGH - DISCHARGE END - TEST NO. 2
FIGURE 44 - TROUGH - INPUT ARRANGEMENT - TEST NO. 2
-76-
-------
The flow test was Initiated and monitored to obtain flow rate
and water samples. Fluid levels were monitored from the third day of the
test, (see Table 24 - Appendix). It was found that Compartment No. 1,
containing Harrold #12 limestone, plugged and overflowed as in Test
No. 1. While maintaining a 7 gpm input rate, the discharge rate de-
clined steadily throughout the initial 30 days of monitoring to 0.5 gpm.
Compartment No. 3, identical in construction to No. 1, con-
tained Harrold #7 limestone. After 5 weeks, the discharge rate was
still equaling the input rate in this compartment, although the water
level in the gauging pipes had risen considerably at every point. How-
ever, the discharge rate began to decline the following week. Compart-
ment No. 2, also containing #7 limestone but having the water introduced
through distribution pipes, followed a similar decline in discharge to
Compartment No. 3.
An extension was added to the sump-of Compartment No. 1 to pro-
vide for an increased hydrostatic head on this roofed portion containing
#12 limestone. The sump was 8 feet high. A 2-foot head was placed on
the compartment and flow started again. This head was maintained and a
discharge of 3 gpm recorded on the following day. The discharge con-
tinued to decrease. Raising the head to 3 feet did not increase the
discharge. It is believed that the increased initial discharge was due
primarily to breaking the seal in the aggregate pores by the construction
of the sump extension.
After 3-1/2 months of monitoring the flow, the test in Compart-
ment No. 2 was discontinued since the flow rate was then down to 0.4 gpm
with 7 gpm input. A sump extension was built on Compartment No. 3 similar
to that already in use on Compartment No. 1. Tests with heads up to 4
feet did not show any increased flow through the limestone. (Table 25 -
Appendix).
V
The water analyses (Table 26 - Appendix) indicated that the
water passing through the #12 limestone with an initial pH of 3.1 neu-
tralized to a pH of about 5.8. The water flowing through the #7 lime-
stone remains essentially unchanged, probably due to less retention
time.
It was then determined to conduct permeable plug Test No. 3.
This test was the placement of #12 limestone in an actual mine to form
a self-sealing plug. Prior to actual field placement, a full-size
laboratory test was conducted to build a plug in a simulated mine
passage. This mine passage was of plywood construction with a section
10 feet wide by 6 feet tall, and a length of 16 feet.
The #12 limestone was conveyed pneumatically from a 220-cubic-
foot pressure tank through 60 feet of 4-inch hose and 20 feet of 4-inch
steel pipe. A 45-degree deflector plate at the discharge end of the
pipe was used to guide proper placement of the limestone, as can be
-77-
-------
seen in Figure 45. This feature permitted the operator to control the
building of the plug. Material was conveyed into the tunnel at the rate
of about 835 pounds per minute, but placement, including loading time,
reduced the overall rate to about 550 pounds per minute. The finished
plug was about 20 feet at the base with a one-foot closure at the top
side. It was not possible on this test to obtain a tight connection be-
tween the roof and aggregate. The plug had an angle of repose of approx-
imately 45 degrees on the front side and 30 degrees on the back side.
About 33 tons of rock were used in building the plug.
Arrangements were then made to build a plug in a mine passage.
A drift opening in Mine No. 62-008 near Clarksburg, West Virginia, was
selected for the test site. This opening was the left-hand Pittsburgh
portal, designated as Opening No. 3. It may be seen in Figure 46 and is
also denoted on the schematic drawing shown as Figure 34. A roof fall
was evident about 100 feet back into the mine drift. A buildup in the
floor about 32 inches high was noted across the drift in the vicinity of
the roof fall. These are evident in Figure 47. This picture also shows
the numerous posts existing in the mine area where the bulkhead was built.
A trailer having two 250-cubic-foot pressure tanks was used to
contain the #12 limestone. The rock was loaded into the tanks from a
dump truck using a belt conveyor system. Both tanks were loaded prior to
the job, then as one tank was emptied, it was refilled while the material
was being placed from the other tank. This allowed semi-continuous opera-
tion, stopping only briefly to move connections from one tank to the other.
The #12 limestone was brought from the quarry in dump trucks
with the tail gate altered to provide an 18-inch square opening and chute
to discharge on the belt conveyor. The total weight of limestone used
was 134,200 pounds.
The first half of the stone delivered was freshly crushed and
dry. This material moved well pneumatically through the 80 feet of 4-inch
hose and 30 feet of 4-inch aluminum pipe used in placement. However, lime-
stone dust in the mine portal was very bad and hindered proper observation
of the work.
The second half of the aggregate was obtained from stockpiles due
to a breakdown of the crusher. This limestone was saturated with water be-
cause of recent rains and did not cause any dust problems, but it was
difficult to move pneumatically because of build-up in lines and valves
of the fine limestone particles in the aggregate.
Material was moved from the tanks using a maximum pressure of
35 psi, assisted at times by an air vibrator attached to each tank. A
special air manifold on the tank, discussed in a previous section, was
used to facilitate movement of the aggregate from the tanks.
Figure 48 shows the seal with about one-half the limestone in
place. The placement pipe with deflector on the end, as well as the drain
-78-
-------
line, can be seen. The angle of repose is about the same as that of the
simulated mine test, or about 45 degrees.
When stoppages were not encountered, a tank of about 17,000
pounds could be emptied in an average time of 20 minutes. This is a
rate of 850 pounds per minute.
The finished seal of aggregate filled the 52-inch by 12-foot
drift with 25 feet of roof contact and 36 feet of length on the base.
It was observed that some settling occurred after several days.
Monitoring of the water flowing through the limestone plug
was begun immediately. The flow data and water analyses for this mine
drift are attached (Table 27 - Appendix). It can be noted that the
water was neutralized from a pH of 3.0 to a pH of 6.3 to 6.9 with a
corresponding change in acidity from 300 mg/1 to an average of 112 mg/1.
The drainage of 3 gpm from the drift did not show any major decline
during the four months of monitoring. By comparison, a flow of 7 gpm
through the same aggregate in the test trough decreased to zero in
55 days. However, the aggregate volume in the test trough was only
12% of that in the drift, while the flow rate was more than double
that through the larger plug. This should indicate that a much longer
time would be required to form enough precipitate to close the pores
in the larger plug installation.
An open fanway about 100 feet north of Opening No. 3 is drain-
ing and is interconnected to the three sealed openings. This fact may
limit the head behind the seals to the approximately 50 inches shown
in the table.
The cost for placing the graded aggregate seal in the mine
drift was $3,048.00. This cost included $756.00 for site preparation,
$237.00 for materials and $2,055.00 for equipment and operators.
-79-
-------
FIGURE 45 - PLACING AGGREGATE IN SIMULATED MINE OPENINGS
FIGURE 46 - DRIFT MINE FOR PERMEABLE PLUG INSTALLATION
-80-
-------
FIGURE 47 - MINE INTERIOR - PERMEABLE PLUG SITE
FIGURE 48 - PLACING LIMESTONE SEAL IN MINE DRIFT
-81-
-------
SEALING OF HIGH-FLOW MINE
The purpose of this phase of the project was to develop methods
whereby remote grouting application could be used in mine drifts having
high water discharge rates. The techniques and materials which were
developed through research in another phase of this project were applied
to the conditions found in the high-flow mine.
Mine No. RT5-2, near Coal ton, West Virginia in the El kins pro-
ject area, was selected for the field tests with the cooperation of the
personnel at the Norton FWPCA laboratory. The location of this mine is
shown on the map attached as Figure 49. A mine map of the Coal ton No. 3
Mine in the Coal ton-Norton Coal Field is attached as Figure 50. The open-
ing selected for field tests is designated on the mine map as Opening A.
Drift mining began in this field in 1894 when the Coal ton No. 1
Mine was started by an individual using pick and shovel mining methods.
Four years later it was taken over by the Davis Coal and Coke Company,
who constructed a large power plant for the mine and 500 coke ovens during
1900. Hand mining continued using short wall machines and, later on, arc
wall machines. All haulage to the main haulways was accomplished by horses,
The first recorded production was 220,040 tons of coal and 87,907 tons of
coke in 1908. The maximum production of this mine was 269,017 tons of
coal in 1915 just prior to ownership changing to the West Virginia Coal
and Coke Company. Coal production declined to 2,769 tons in 1937.
The Coal ton No. 2 Mine was started in 1900 and abandoned in
1928. The Coalton No. 3 Mine, in which the field tests occurred, began
in 1904 to 1905 and operations ceased in'1923. All of these Coalton
Mines were hand mined and pillared during retreat mining. The remainder
of the coal in Coalton No. 3 Mine was mined up to 1950 and was hauled
out of the Norton No. 2 Mine.
Norton No. 1 Mine was started in 1902 by Davis Coal and Coke
Company who also owned the Coalton Mines. The first recorded coal pro-
duction at Norton was 255,792 tons in 1918. Steel pan conveyors were
installed in 1920 in the Norton No. 2 Mine. This was supposedly the
first mechanized mine in the United States. The maximum production of
the Norton Mine was 540,403 tons in 1942.
There are presently two active mining companies in the area.
The Demott Coal Company, which started in 1952, and Roaring Run Coal
Company are operating in the Norton No. 2 Mine. Roaring Run Coal
Company produces approximately 400 to 500 tons per day. During the
1940's the entire area was strip mined.
The seam mined in all these mines is in the Allegheny series
and is called the Kittanning Seam. A geologic section of Mine No. RT5-2
is included in the Appendix as Figure 66 to show the formations which
occur in the area. This mine is in the El kins project area where
-83-
-------
- ( r-v <&•} { , '
^«S
HI
•: " v
9 *1 >
,V'V MxiH.n ! ,.^ -V^i
^ c'-" ' Oi ,;¥
^, ^c
'is
,-a Ivvrl-
FIGURE 49 - LOCATION MAP - MINE NO. RT5-2
-84-
-------
considerable reclamation work was accomplished in 1966 and 1967 by the
Federal Water Pollution Control Administration.
The field testing was accomplished in the following manner:
The entire flow was diverted through a drain line placed on the floor
of the drift. A rear bulkhead of quick-setting, self-supporting
cementitious material was constructed in front of the wet seal. The
drift was filled with aggregate after pipes were placed for a subse-
quent grouting operation. A front bulkhead of the quick-setting mate-
rial was then placed across the front of the aggregate. The aggregate
was grouted through the pipes placed for that purpose.
When the head of water impounded behind the seal increased,
the pressure caused water to leak out from an unknown adjacent opening.
Remedial work on the second opening was accomplished by constructing a
bulkhead of graded aggregate and agricultural lime in the drift to seal
the opening.
The two openings which were sealed were at one end of the
Coalton No. 3 Mine, far from the main heading. These openings were
originally driven out for use as air supply openings. Later these were
used as entries. Since they were close to the stables in the town of
Coalton, the horses were moved into the mine through one of these open-
ings, and the other was used for a fanway. Opening No. 1 of Mine No.
RT5-2, which was selected for the field test, was considered suitable
because the site was easily accessible and provided sufficient room
for the necessary equipment to be used in placing the mine seal. It
also had a sound roof and good timbering and a water flow in excess of
50 gallons per minute. The coal beds have a slight dip and flooding of
the entire section of the mine was thought to be possible. In addition
to these qualifications, the natural drainage patterns surrounding this
mine would tend to prevent the flooding of any populated area in case
of leakage or failure of any opening in the mine. Other openings were
known to exist and many of these were sealed only with a clay seal. A
view of this opening prior to the field test is shown in Figure 51.
During the El kins project, Opening No. 1 had been closed by
placing a wet seal in the main shaft and a dry seal in the crosscut
coming into the main passage from the right side. These are shown in
Figure 52, which is a drawing of the remedial construction accomplished
during the field tests. The wet seal consisted of a concrete footing
and a double course of concrete blocks completely filling the mine
drift. Openings had been provided through this wall by omitting two
8 by 16-inch blocks on each side of the mine passage, 16 inches above
the floor of the mine. About 4 feet in front of this wall was a dam
of concrete blocks which retained about 3 feet of water. This prevented
air entry into the mine and permitted measurement of the water flow.
The dry seal was a solid concrete block wall placed in a crosscut to
the right of the passage about 4 feet back into the crosscut. To pre-
pare this opening for placement of the mine seal, the dam in front of
the wet seal was removed and a portion of the wet seal, including the
-85-
-------
FIGURE 50 - MINE MAP - COALTON NO. 3 MINE (RT5-2)
-86-
-------
.
:.
•-•-
FIGURE 51 - VIEW OF MINE NO. RT5-2
-87-
-------
I
. 0
CO
XISTING CONCRETE BLOCK
WET SEAL
EXISTING CONCRETE BLOCK
DRY SEAL
REAR CEMENT
BULKHEAD
EXISTING
CONCRETE
BLOCK 30'
DRY SEAL
ROOF \ —-r^r^r—- r
LIMESTONE -
I AGGREGATE!
•v,-. -.-r.--::..;^'
viV GROUTED''--1
:••/AGGREGATEj
8 HOLE USED FOR
INITIAL REMOTE T-V
"5 INSPECTION
FRONT CEMENT
BULKHEAD
SEAL*2 /
EXISTING CLAY SEAL
> DIRECTION OF OLD CUT
SEAL
PLAN VIEW OF REMEDIAL CONSTRUCTION - MINE NO. RT5-2
FIGURE 52
-------
footing for the dam and wall, was broken out to permit the installation
of a drain line on the floor of the mine. To achieve a valid test,
additional openings were made through the wet seal, allowing the mine
water direct access to the new seal as the mine is flooded.
It was not necessary to remove the concrete block wet seal
wall since there was sufficient room between this wall and the cross-
cut to obtain a seal on the drift wall, roof and floor with the rear
bulkhead. The rear bulkhead was built in front of the wet seal wall
and was self-supporting.
Additional preparation included thoroughly washing the floor
and walls of the drift with a high-volume pump to aid in the bonding of
the cement material to the wall and floor of the mine.
A drain line of 4-1/2-inch OD Fibercast pipe, 40 feet in
length, was installed and extended 5 feet past the concrete block wall
of the wet seal. This line was placed on a 6-inch bed of AASHO No. 67
crushed limestone which raised the pipe above the floor of the mine to
insure a seal around the pipe. A screen analysis of this limestone is
included in the Appendix in Table 28. The drain pipe was also covered
with Halliburton Casing Kote. This is a bonded layer of angular sand
grains around the pipe to give better hydraulic sealing characteristics
to the drain pipe surface.
The mine seal was constructed in four stages. The first
stage consisted of constructing a rear bulkhead of quick-setting, self-
supporting cementitious material just in front of the concrete block
wall of the wet seal.
The second stage consisted of the placement of six 2-inch
grout pipes and then filling the area in front of the rear bulkhead
and into the crosscut to the dry seal with AASHO No. 67 crushed lime-
stone.
The third stage was the construction of the front bulkhead of
the same material as the rear bulkhead. This bulkhead was built against
the slope of the limestone aggregate to contain the grouting fluid used
in the aggregate consolidation.
The final stage was the consolidation of the limestone aggre-
gate with a grouting fluid to make the material impervious between the
two bulkheads.
The process and materials used to build the rear and front
bulkheads were the same as used to build three previous bulkheads in
Mine No. 62-008. This work was recorded in a previous section of this
report. In brief, this process involves preparing two slurries sep-
arately and mixing them together as they are being pumped into the
mine. The slurries react to give a viscous quick-setting material
which is able to support its own weight as it builds.
-89-
-------
The pumping and mixing were accomplished using the pumps on a
mobile pumping unit. As the materials were mixed, each slurry was placed
in one of the two trailer-mounted pressure tanks.
From these tanks, the slurries were pumped simultaneously through
separate lines to a common point at a wye connection. The resultant slurry
then passed through a special mixing section of line as the reaction con-
tinued. This slurry was discharged through a nozzle designed for this par-
ticular application. The piping and mixing sections were mounted on a two-
wheel dolly so the placement of the material could be controlled by an
operator standing at or near the entrance to the mine.
The bulkhead was then built as previously described. It con-
tained approximately 14 cubic yards of sodium silicate cement slurry.
The angle of repose on the front and rear sides of the bulkhead was
approximately 55 to 60 degrees.
After placement of the rear bulkhead, the slurry tanks used were
cleaned to prepare them for containing the AASHO No. 67 aggregate (West
Virginia No. 7) which was to be placed pneumatically into the mine. The
tanks were loaded from dump trucks with a belt conveyor.
Simultaneous to the preparations and loading of the aggregate,
grout pipes were installed in the mine drift in order to permit grouting
of the limestone. All pipes were covered with Halliburton Casing Kote at
the contact area through the front bulkhead. Three pipes of PVC plastic
were installed on each side of the drift, one near the floor, one at the
roof timbers and one about midway between the two. The pipes were nominal
2-inch size fitted with 2-inch plastic ball valves on the outer end. The
pipes at the roof were 10.5 feet long. The center pipes were approximately
20 feet, and the lower pipe was 15 feet on the left-hand side and suffi-
ciently long on the right-hand side to go back to the crosscut and bend
into the crosscut near the dry seal. An extra pipe to be used as a hydro-
static measure was installed at the center of the opening along the roof
line. It was not fitted with a valve, but had a 90-degree ell and a riser
pipe for measuring the head during the grouting operation. The grout pipe
arrangement is shown in Figure 53.
The dolly to be used for placement of the limestone in the mine
was basically the same as that used for building the cement bulkhead,
except that the 2-inch mixing pipe from that operation was replaced with
a 4-1/2-inch OD aluminum pipe with a special 45-degree deflector on the
end of the pipe.
The pneumatic placement of this limestone was accomplished by
placing air on the pressure tanks containing the aggregate and using a
special air manifold on the bottom of the tanks in the same manner as
described previously. The material was moved with air through a 4-inch
hose and the pipe on the dolly into proper placement in the mi ne drift.
During this operation, 92,200 pounds of AASHO No. 67 aggregate were placed
in the mine drift. This conveying was done in 148 minutes for an average
-90-
-------
rate of 622 pounds per minute while operating. The overall time for
preparation, placing of the grouting pipes, loading, conveying and
demobilizing was 11-1/2 hours.
The next stage was the construction of a quick-setting cement
bulkhead on the front slope of the aggregate which had been conveyed
into the mine drift. Figure 54 shows the beginning of this operation.
The materials used here were the same as those used in building the
rear bulkhead. They were prepared and placed in a similar manner.
Figure 55 shows the completed front bulkhead. About 9 cubic yards of
slurry were used. The grout pipes with valves attached can be seen
on either side. The hydrostatic pipe is evident coming from the top
center portion of the bulkhead. The drain pipe can be seen on the
floor of the mine in the lower left side.
On the following day, the grouting operation was performed
using the same pumping equipment which had been used in building the
bulkheads. The grouting material used was Halliburton LIGHT Cement.
This material was a mixture of 65% portland cement and 35% fly ash,
6% bentonite and 9.9 gallons of water per bulk cubic foot of material,
or 61 parts cement, 26 parts fly ash, 5.2 parts bentonite and 82.5
parts water by weight. This resulted in a slurry which weighed 12.7
pounds per gallon. The grouting slurry was pumped into the lower line
on the right and left side simultaneously, at an average injection
pressure of 10 psi. The valves on the upper lines were left open until
the grout slurry flowed from them, at which time they were closed and
the grouting was continued. The grout injection was moved progressively
to higher grout pipes as the work proceeded.
Pumping was continuous until a fill-up was obtained to within
a few inches of the roof. At that time, leaks appeared around the top
corners of the front bulkhead, so the slurry was altered by adding
shredded cellophane, sawdust and shredded cane fiber. This material
successfully closed the leaks at the top of the bulkhead and the grout-
ing was concluded at a final pressure of 15 psi.
The building of the seal in the drift mine, consisting of the
rear bulkhead, placement of the grouting pipes and the aggregate, and
the building of the front bulkhead, required two days. The grouting
required about half of one day. The preparation had required several
days prior to this. The completed seal was approximately 25 feet in
length and is shown in detail on the drawing attached as Figure 56.
The total cost for this seal, including site preparation cost of
$1,079.00, was $9,463.00 for materials and equipment.
A valve was installed on the drain line and closed 48 hours
after the final grouting to commence flooding of the mine. Monitoring
of the fluid level behind the seal then began (Table 29 - Appendix).
At the time of closing the valve, the flow rate coming from the mine
was 58.5 gallons per minute.
-91-
-------
TOP
34' II"-
/
_EL
10'-
17'-
REAR
fBULKHEAD
IZE
13'
GROUT PIPE ARRANGEMENT
SEAL NO. I - MINE NO. RT5-2
FIGURE 53
-92-
-------
FIGURE 54 - BEGINNING OF FRONT BULKHEAD CONSTRUCTION
-93-
-------
_ .
•tr
FIGURE 55 - COMPLETED FRONT BULKHEAD - MINE NO. RT5-2
-94-
-------
PLAN
7
mw~\
* • - * . X
FRONT
BULKHEAD/^. .-• •••••. •/
'/GROUTED .',?
'.'• AGGREGATE
REAR
;/BULK HEAD
SECTION A-A
CONSTRUCTION DETAIL
SEAL NO. I - MINE NO. RT5-2
FIGURE 56
-95-
-------
On the sixth day after monitoring began, the head was 2.92 feet
behind the seal. A small leakage was noticed coming through the backfill
at the right side of the drift. The following day the head was 3.22 feet
and the leak had increased substantially. The water was tested and found
to have a pH of 2.8, indicating it was mine drainage.
Hand digging into the back slope right of the portal confirmed
the direction of the water source. Additional excavation with a backhoe
disclosed an opening about 6 feet to the right of the main portal and the
remains of a masonry seal. A view of this excavation is shown in Figure 57.
The drain valve was then reopened to prevent further head buildup.
It was decided to attempt to drill into the suspected opening
and use closed circuit television to inspect it. Accordingly, a space
was cleared for positioning an augering machine. An 8-inch hole was
drilled in the top part of the coal vein as shown on Figure 52. At a
depth of 18 feet, a void was encountered. A length of 7.5-inch OD casing
was shoved into the hole to prepare for use of a television camera.
A television camera was pushed into the casing until the front
of the camera was flush with the void. An attempt was made to identify
the void which had been encountered through the picture received from
the television camera. It was very difficult to determine exactly what
was being shown on the screen. It appeared that either the lighting was
inadequate or the scope of the camera was limited to the portion imme-
diately in front of the camera. In looking at the dimly lighted picture
on the television screen, something could be seen which appeared to be a
pile of debris or a roof fall. Other than this identification, it was
not possible to determine any other usable information.
It was then decided to continue excavation to remove additional
material and completely expose the opening that had been encountered. This
opening, designated as Opening No. 2, was found to be 28 feet to the right
of Opening No. 1. After cleaning the front side, the opening was deter-
mined to be 12 feet wide. It was filled to the top of the coal with debris
from the roof fall, leaving only 2 to 4 feet of void space between the top
of the debris and the new roof. The existing roof was of a fossiliferous
shale and appeared to be in good condition. A view of this exposed open-
ing is shown in Figure 58.
After looking at the opening as it now appeared, it could be
seen that the view given on the television camera was accurate and that
the camera was located so that it was looking across the top of a large
roof fall. The right-hand wall of this opening was diagonally about 15
feet in front of the camera. The existing mine roof after the fall was
approximately 4 feet above the horizontal line of the camera for the
entire distance to the opposite wall. Without seeing the exposed open-
ing in full, it was difficult to interpret the information shown by the
television camera. It is assumed that this method of viewing mine drifts
could be useful, but only after additional work was done to make the
lighting system and lens adaptable to this type service.
-96-
-------
Since the opening was large and the roof fall continuous along
the drift, a backhoe was used to remove the roof fall. As the clean-
out progressed, protective timbering was installed using posts of 5-inch
by 7-inch green oak timber on each side and cross members at the roof.
Five of these were spaced 2 to 4 feet apart and extended into the open-
ing for 12 feet.
As the cleaning progressed, water flow from Opening No. 2 in-
creased with a corresponding reduction of the flow from the drain line
of Opening No. 1. The opening was cleaned to a distance of 30 feet from
the face. Very large boulders were encountered and had to be dragged
out by chain.
Since the floor of the mine was lower than the ground outside
the mine, a pump was used to keep the water pumped from the mine. The
water flow increased as the clean-out progressed, so a second pump was
installed. When the opening was sufficiently cleaned and the water
level lowered, it was found that the main flow of water came from a
point 20 feet back in the opening at the center between the two walls.
Digging and probing disclosed a 15-inch ID drain line of vitrified clay
laid in a ditch cut in the floor of the mine. The ditch, about 3 feet
wide by 2 feet deep, was cleaned out by hand to the front of the mine.
Most of the water coming from Opening No. 2 was coming through this
drain tile which continued back into the mine under the roof fall for
an unknown distance.
A plug was fabricated for the 15-inch drain tile from discs
of 3/4-inch plywood and 3/8-inch gasket material, spaced 18 inches apart
with 2 by 4-inch wood spacers. A 4-inch plastic pipe was placed through
the plug. This pipe was then extended outside the mine so that it would
serve as a drain line. The plug was inserted into the drain tile causing
the flow to be diverted into the 4-inch drain line and out of the mine.
The special plug was grouted into the 15-inch drain tile by
filling the ditch with cement grout slurry. The 4-inch drain line was
brought out of the ditch on a slope so that the outer end was exposed
outside the mine.
It was decided to place a limestone aggregate seal in the
second opening. This seal was to be grouted along the top to help in-
sure a seal across the roof section. It was suggested that agricultural
lime be blended with the aggregate, if preliminary tests indicated that
this was practical, in order to reduce the permeability for longer re-
tention time and give increased surface area to cause faster neutral-
ization.
A test was made at the laboratory facilities to observe the
results of conveying a blended aggregate-lime mixture. A total of
10,000 pounds of 3/8-inch limestone chips containing approximately 15%
agricultural lime was conveyed pneumatically from a 220-cubic foot
-97-
-------
pressure tank through 54 feet of 4-inch line. An initial pressure of
10 psi was placed on the tank containing the aggregate using a 110 scfm
compressor. Then air from a 428 SCFM compressor was introduced into the
bottom discharge manifold of the tank at a pressure of 30 psi. The aggre-
gate moved steadily through the line and was discharged in an open area
against a board wall. Extreme dusting was encountered which would be a
problem when placing the material in a mine drift. A side view of the
aggregate being discharged, shown in Figure 59, emphasizes the dust
problem.
An excellent rate of 1,207 pounds per minute was obtained for
the single tank load tested, but some means of alleviating the dust
would be desirable for field application.
Preparation was made to use this type aggregate blend for a
seal installation in Opening No. 2. Grout lines of 2-inch plastic pipe
were installed in the opening as shown in Figure 60. The purpose was
to grout only the upper part of the aggregate to insure a seal against
the roof of the mine in the event any settling should take place. Pipes
were placed along both sides of the opening at the roof timbers, in the
center of the vertical opening and on the floor.
The opening was filled with a large quantity of AASHO No. 8
crushed limestone blended with approximately 15% agricultural lime.
Screen analysis for AASHO No. 8 (West Virginia No. 12) aggregate is in
Table 28 of the Appendix. The aggregate seal was placed against a roof
fall in the drift about 30 feet from the opening at the highwall. The
agricultural lime consisted primarily of the fine limestone dust that
is obtained when limestone is crushed. The limestone was stockpiled on
the ground and then loaded into a 500 cubic foot pneumatic trailer using
a highlift and belt conveyor. The agricultural lime was in 50-pound
bags and was blended with the crushed limestone on the belt conveyor
as it was being put into the pneumatic tank.
The rock and lime mixture was conveyed from the pneumatic
tank into the mine drift through 40 feet of 4-inch rubber hose and the
20 feet of 4-inch pipe contained on the placement dolly. This was the
same equipment used for placement of the rock in Opening No. 1.
One problem which was prevalent in this test was the extreme
amount of dust created when the limestone was blown into the mine drift.
In order to combat this problem, the ratio of lime dust was reduced to
10% and a pump was used to spray a one-half inch jet of water against
the deflector plate on the discharge end of the dolly at the same time
as the limestone was impinging on the plate. This helped to some degree,
but even using this water spray the dust problem was still bad.
A total of 330,000 pounds of material was pneumatically placed
in the mine in a net time of 346 minutes, for an average rate of 955
pounds per minute. The overall time for the placement was 23 hours,
including all the downtime, loading time, etc., involved in the operation.
-98-
-------
FIGURE 57 - OPENING EXCAVATED RIGHT OF MAIN PORTAL - MINE NO. RT5-2
-99-
-------
FIGURE 58 - OPENING NO. 2 AFTER EXPOSING FRONT - MINE NO. RT5-2
-------
I
o
'
FIGURE 59 - DUST CREATED WITH PNEUMATIC CONVEYING OF AGGREGATE CONTAINING AGRICULTURAL LIME
-------
O
ro
i
H±I(L±L
-Y
i=m
{H'6" LONG)
(20* LONG)
NOTE'GROUT PIPES *2AND*4 ARE PERFORATED WITH j"
HOLES EVERY 3 FEET AND OPEN ENDED. GROUT
PIPES *5 AND *6 ARE PERFORATED EACH FOOT ON
PORTION ACROSS OPENING ONLY. GROUT PIPES
*.T AND*8 HAVE y" PERFORATIONS EACH FOOT ON
INNER 5 FEET WITH END PLUGGED.
•12'
?
li!
GROUT PIPE PLACEMENT-OPENING NO. 2
FIGURE 60
-------
This work was spaced over 2 days' working time.
After the material was all placed in the drift opening, a valve
was installed on the 4-inch drain line and a vertical length of clear
tubing was provided.to measure the water head behind the limestone in the
mine.
The grouting of the upper portion of the limestone plug was
accomplished by pumping a Halliburton LIGHT Cement slurry as a grouting
slurry. This was the same slurry that was used to grout the aggregate
in Opening No. 1 of this mine. The grout was pumped into the two upper
pipes on either side of the opening. The grouting was accomplished by
pumping approximately 100 cubic feet of grout slurry into the upper
portion of the limestone.
Figure 61 shows a section view of the seal which was installed
in Opening No. 2 of Mine RT5-2. The aggregate seal is approximately 30
feet in length.
Since the area immediately in front of the mine was low and
filled readily with water that came through the limestone, a retain-
ing wall was built in front of the opening with a 90-degree weir in-
stalled to measure the flow rate. A view of the completed installation
at Opening No. 2 is shown in Figure 62.
The graded aggregate seal placed in Opening No. 2 was con-
structed at a cost of $8,463.00. This cost included $3,447.00 for site
preparation, $1,696.00 for materials and $3,320.00 for equipment and
operators. This cost was higher than a previous seal of this type
because of the excessive excavation required to prepare the opening,
the extra materials required for grouting the upper section and a
corresponding increase in the necessary equipment.
After the seal was placed, monitoring was resumed for Opening
No. 1 and initiated for Opening No. 2. The head of water impounded and
the flow through the aggregate seal of Opening No. 2 was recorded (Table
30 - Appendix). The initial flow of 25 gpm with a head of 3.3 feet of
impounded water diminished to a flow of 3.6 gpm with a head of 5.44 feet
in about 24 days. Since the field work on this project ceased about 30
days after the seal was installed, the monitoring data stopped at this
point. However, monitoring was continued by Federal Water Pollution
Control Administration personnel from the Norton, West Virginia Acid
Mine Drainage Laboratory.
The water impounded behind the seals in the two openings had
a pH of 2.9 to 3.2, an acidity of 512 to 765 mg/1, an iron content of
104 to 282 mg/1 and sulfate of 868 to 1,152 mg/1. After passing through
the limestone seal, the water neutralized to a pH of 5.6 to 6.3, an
acidity of 0 to 103 mg/1, iron content of 36 to 57 mg/1, and sulfate of
780 to 940 mg/1. The hardness did increase as a result of the lime
treatment, however (Table 31 - Appendix).
-103-
-------
It was impossible to tell whether the leakage through the lime-
stone seal was channeled through one area or across the entire seal because
of the condition in front of the opening. This area was low and water re-
mained impounded, covering the lower portion of the seal and obscuring
observation of the leakage.
Additional information on the monitoring of the two openings
after cessation of field work on the project has been furnished by the
Norton FWPCA personnel. The head on the impounded water behind both seals
has slowly increased, reaching 6.48 feet in Opening No. 2 and 4.74 feet at
Opening No. 1 by December 30, 1969. The flow remains at 3.6 gpm through
the limestone seal. Water quality likewise is similar to that reported in
Table 31.
-104-
-------
o
en
i
10'
HIGHWALL
-*
l—10 .
\V • '. . ' ''.'/.'•' ' , '
U ' ' s ^-30 AVG. -
A7-.'•'• :/•.'•'.•.•• .;•••.•
si'/,"- •••.-'•••:•.•.. :.'•.•
H- •••'::•::.•.-•.. •.-:••.:,
::':'VX
. ;.' . •' •". •/•/'. ..". •; -. -;.\
'•'•'.'''•"' • '• •'•'• »••*•'•'•/".:.•• / :>
- • '• • • •'.'/ •• '•'. ''.'•^-'":~i~'~-^'—'r-
"
20.5
23.3-
SECTION VIEW OF SEAL - OPENING NO. 2
FIGURE 61
-------
o
cr,
FIGURE 62 - COMPLETED INSTALLATION - OPENING NO. 2
-------
ACKNOWLEDGEMENTS
In the development and accomplishment of the studies and in-
vestigation requisite to the preparation of this report, the valuable
assistance and suggestions of the Project Officers are gratefully
acknowledged. During the Contract, four different Project Officers
were assigned. They are listed below in chronological order:
Dr. Edward J. Martin - Washington, D.C.
July 15, 1968 to August 27, 1968
Mr. John R. Hyland - Wheeling, West Virginia
August 28, 1968 to March 16, 1969
Dr. Donald L. Warner - Cincinnati, Ohio
March 17, 1969 to September 17, 1969
Mr. Donald J. 0'Bryan - Washington, D.C.
September 18, 1969 to end of Contract
The analyses of the many water samples by the Federal Water
Pollution Control Administration Laboratory in Wheeling, West Virginia
were deeply appreciated.
Mr. Paul W. Hornor, Consulting Mining Engineer of Clarksburg,
West Virginia, rendered valuable assistance in general consultation
during the contract.
The cooperation of the personnel from the Federal Water
Pollution Control Administration Laboratory at Norton, West Virginia
is also gratefully acknowledged.
The helpful suggestions and comments from Mr. Ronald D. Hill,
Robert A. Taft Water Research Center, Federal Water Pollution Control
Administration, Cincinnati, Ohio were greatly appreciated.
-107-
-------
REFERENCES
American Petroleum Institute, Publication RP 13B,
November, 1962, 15p.
Halliburton Company, Part II, "Selection and Recommendation
of Twenty Mine Sites", August 23, 1967, 161p.
Halliburton Company, Part IV - "Additional Laboratory and
Field Tests for Evaluating and Improving Methods for
Abating Mine Drainage Pollution", May 21, 1968, 134p.
-109-
-------
PATENTS
Two patent applications have resulted from the work con-
ducted on this contract. The description of these patents is given
bel ow:
(1) FWP-1531
E. L. Paramore et al
Filed: May 13, 1969
Serial No.: 824,159
For: "Forming Self-Supporting Barriers
in Mine Passages and the Like"
(2) FWP-1564
Tommy R. Gardner et al
Filed: July 9, 1969
Serial No.: 840,494
For: "Method of Plugging Mine Passages
Having Water Emanating Therefrom"
-m-
-------
ABBREVIATIONS
alka.
AASHO
bpm
cfs
cm
Cr.
disch
ft
gpm
hr
hrs
ID
in.
Ibs/ft
Ibs/gal
Ibs/min
mg/1
ml
No., #
OD
psi
scfm
US6S
alkalinity
American Association of State Highway Officials
barrels per minute (1 barrel = 42 gallons)
cubic feet per second
centimeters
Creek
discharge
foot, feet
gallons per minute
hour
hours
inside diameter
inch, inches
pounds per foot
pounds per gallon
pounds per minute
mi 11 i grams 'per 1 i ter
mi 11i1iters
number
outside diameter
pounds per square inch
standard cubic foot per minute
United States Geological Survey
-113-
-------
APPENDIX
-------
10
15
10
SANDSTONE
SANDY
SHALE
SHALE
SHALE
a
SLATE
PITTSBURGH
COAL
FIGURE 63-GEOLOGIC SECTION-MINE NO. 14-042A
18-20'
SANDSTONE
SHALE
REDSTONE
COAL
LIMESTONE
SHALE
PITTSBURGH
COAL
FIGURE64-GEOLOGIC SECTION-MINE NO.40-016
-117-
-------
SHALE
to'i
ID'
REDSTONE
COAL
SHALE
SANDSTONE
LIMESTONE
SHALE
SLATE
PITTSBURGH
COAL
FIGURE 65-GEOLOGIC SECTION-MINE NO. 62-008
20'-
• • • •
20'
I5't
5't
• • • • •
'
•-•••*.
SHALEY
LIMESTONE
SILTSTONE
(WITH SANDY STREAKS)
SILTY
SANDSTONE
SANDY
SHALE
CARBONIFEROUS
SHALE
LOWER KITTANNING
COAL
FIGURE 66-GEOLOGIC SECTION-MINE NO. RT5-2
-118-
-------
TABLE 1
FLUID PROPERTY TEST1- MINE NO. 14-042A
Gelled Fluid
Batch Batch Size
No. (gallons)
Viscosity
Marsh Funnel
(seconds)
Filter Los_s2_
Filter Cake
cc (inches)
Time After
Mixing
(hrs & mins)
1
1
2
2
2
3
3
_3
4
4
5
5
820
820
10,000
10,000
10,000
10,300
10,300
_3
10,300
10,300
10,300
10,300
39
43
47
58
63
56
55
62
48
47
45
48
19.1
17.0
18.2
17.2
16.8
16.6
16.2
15.8
15.8
15.8
16.4
17.1
2/32
3/32
2/32
2/32
2/32
3/32
3/32
2/32
2/32
2/32
2/32
12/32
:05
1:50
:05
1:00
1:30
:20
1:10
-
:25
1:10
:30
22:30
1 Tests were conducted using standard API test procedures given
in API Bulletin RP 13B.
2 Filter loss measured in Baroid Filter Press using filter paper.
Reading shown was measured at 30 minutes with 100 psi pressure.
3 Sample taken from fluid inside mine about 2-1/2 feet below
fluid level.
-119-
-------
TABLE 2
MINE FLUID DATA - MINE NO. 14-042A
Date
9-30-68
9-30-68
9-30-68
10- 1-68
10- 1-68
10- 2-68
10- 2-68
10- 2-68
10- 3-68
10- 3-68
10- 4-68
10- 4-68
10- 5-68
10- 6-68
10- 7-68
10- 8-68
10-10-68
10-12-68
10-17-68
10-29-68
11- 1-68
11- 8-68
Gelled
Fluid Added
Time (gallons)
7:00 a.m.
11:47 a.m. 800
3:25 p.m. 9,400
7:30 a.m.
10:55 a.m. 10,300
7:30 a.m.
10:32 a.m. 10,300
3:45 p.m.
7:00 a.m.
12:45 p.m. 8,200
8:00 a.m.
11:03 a.m. 2,100
7:30 a.m.
10:30 a.m.
7:30 a.m.
10:00 a.m.
3:30 p.m.
3:30 p.m.
-
-
-
— —
Fluid
Level
JM
5.00
5.10
5.52
4.94
5.69
5.19
6.67
6.31
6.15
9.04
7.81
8.98
8.83
8.33
8.17
7.88
7.60
7.40
6.96
6.56
6.54
—
Level
Change
(ft)
-
+0.10
+0.42
-0.58
+0.75
-0.50
+1.48
-0.35
-0.17
+2.90
-1.23
+1.17
-0.15
-0.50
-0.17
-0.29
-0.27
-0.21
-0.44
-0.40
-0.02
—
Discharge
Rate
(gpm)
0.40
0.40
0.50
0.40
0.50
0.50
0.75
0.55
-
-
0.55
0.50
0.55
0.55
0.55
0.45
0.60
0.50
0.54
0.48
-
0.48
Sample
No.
2
-
-
3
4
5
-
-
-
-
6
-
7
8
9
10
11
-
12
13
-
14
-120-
-------
TABLE 3
WATER MONITORING DATA - MINE NO. 14-042A
Sample Fluid Level Flow Rate
Date No. (feet) (gpm)
11-21-68 15 6.40 0.48
H-26-68 16 6.36 0.50
12- 5-68 17 6.30 0.60
12-13-68 18 6.28 0.70
12-18-68 19 6.25 0.52
1- 2-69 20 6.20 0.80
1- 8-69 21 6.23
1- 9-69 22 - 1.05
1-16-69 23 6.20 0.95
1-24-69 24 6.17 1.25
1-30-69 25 6.02 1.80
2- 4-69 26 6.02 3.00
2-10-69 27 6.03 2.40
2-18-69 28 5.96 2.10
2-24-69 29 6.04 2.10
3- 3-69 30 6.01 1.38
3-10-69 31 6.04 1.10
3-17-69 32 6.01 0.95
3-24-69 33 6.00 1.10
4- 1-69 34 6.00 0.85
4- 7-69 35 6.01 1.40
4-14-69 36 - 1.55
4-21-69 37 6.02 1.50
4-28-69 38 6.00 1.10
5- 5-69 39 6.03 1.15
5-12-69 40 6.04 1.70
5-19-69 41 6.04 1.40
5-26-69 42 6.06 1.30
6-2-69 43 6.08 1.10
6-10-69 44 6.08 0.95
6-16-69 45 6.081 0.70
6-23-69 46 6.06 0.65
6-30-69 47 6-. 04 0.60
7- 7-69 48 6.04 0.60
7-21-69 49 6.02 0.65
7-28-69 50 6.00 0.62
8- 4-69 51 6.00 0.52
8-11-69 52 5.98 0.58
8-18-69 - 5.98
8-25-69 - 5.98
9- 2-69 - 5.96
9-22-69 - 5.92 0.55
10- 6-69 - 5.95 0.51
10-13-69 - 5.95 0.40
1 Water-level recorder was still making a straight line after a hard
rain. It was suspected that heavy gel around float was preventing
float to fluctuate as water level varied.
-121
-------
TABLE 4
WATER ANALYSES - MINE NO. 14-042A
ro
r\>
Date
Sampled
9-21-68
9-30-68
10- 1-68
10- 1-68
10- 2-68
10- 4-68
10- 5-68
10- 6-68
10- 7-68
10- 8-68
10-10-68
10-17-68
10-29-68
11- 8-68
11-21-68
11-26-68
12-13-68
12-18-68
12-26-68
1- 2-69
1-16-69
1-24-69
1-30-69
2- 4-69
2-10-69
2-18-69
2-24-69
3- 3-69
Cond.
(umhos/cm)
5966
6100
6366
5720
5834
5410
5480
5580
5370
5460
5240
5360
5460
5540
5660
4750
5440
5442
5280
5220
5200
4600
5020
4850
4400
4800
4808
4880
Acidity
as CaCCh
fiH (mg/D
2.8 4520
3.2 4650
3.5 4510
3.6 4920
3.7 4730
3.0 4480
3.0 4140
3.0 3970
3.0 3985
3.1 4360
3.2 4350
2.9 4340
3.0 4180
2.8 3910
3.2 4030
3.2 4070
2.8 4290
3.0 4120
3.0 3700
2.9 3700
2.9 4200
2.9 3875
2.9 3700
2.9 3670
2.8 3900
2.6 3620
2.7 3640
3.1 3815
Alka.
as CaCXh
(mg/1)
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
Hardness
as CaC03
(mg/1)
1620
1400
1470
1325
1270
1215
1415
1300
1330
1030
990
1175
1400
1400
1250
1370
1065
1160
1570
1480
1285
955
1330
510
890
1630
1360
710
Iron
(mg/1)
1020
876
912
900
960
900
1032
852
831
792
749
852
864
840
936
480
696
840
840
900
816
832
701
692
625
790
685
660
Sulfate
(mg/1)
5850
5460
4316
5590
6110
6110
5330
5330
4810
4940
4550
6500
4550
5200
5850
6110
5070
4680
5850
5850
5590
5050
3900
4550
4345
4400
5100
3600
Alumi num
(mg/1)
340
404
420
334
410
480
880
380
304
394
334
224
304
364
260
132
281
414
180
205
284
131
184
177
185
225
250
115
-------
ro
Co
TABLE 4 - Concluded
WATER ANALYSES - MINE NO. 14-042A
Date
Samp! ed
3-10-69
3-17-69
3-24-69
4- 2-69
4- 7-69
4-14-69
4-21-69
4-28-69
5- 5-69
5-12-69
5-19-69
5-26-69
6- 2-69
6-10-69
6-16-69
6-23-69
6-30-69
7- 7-69
7-21-69
7-28-69
8- 4-69
8-11-69
Cond.
(umhos/cm)
4600
4800
4738
4700
4396
4600
4322
3940
4580
4080
4268
3920
4300
4100
4100
4240
3734
4018
3619
3610
2582
3140
£H
3.0
3.2
2.6
2.9
3.0
2.6
3.1
3.1
1.9
3.3
2.6
2.5
2.8
2.8
3.2
3.1
3.1
3.0
3.0
2.6
2.9
2.6
Acidity
as CaCOo
(mg/ir
3646
3420
3000
3335
2820
2700
2975
2470
3306
2560
2760
2800
2745
2725
2550
2820
2650
2765
1285
2250
2670
2475
Alka.
as CaCOq
(mg/ir
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
Hardness
as CaCOo
(mg/ir
28
1010
575
960
940
830
940
575
915
100
965
1208
965
1020
940
1220
1140
1080
1440
1080
1150
1085
Iron
(mg/1)
735
837
631
653
587
720
646
553
238
522
635
217
498
571
44
475
622
376
413
411
538
476
Sulfate
(mg/1)
4200
4400
3900
4900
4500
4000
4200
3200
3300
3800
4200
4200
4000
4100
3500
4300
3825
3100
3105
2700
2600
3440
Aluminum
(mg/1)
160
164
178
160
175
167
160
145
161
165
158
152
156
146
136
136
151
153
143
126
149
148
-------
TABLE 5
WATER MONITORING DATA - MINE NO. 40-016
Fluid Level - (inches
Date Opening #TOpening #
ro
9-20-68
10- 4-68
10- 7-68
10-17-68
10-21-68
10-24-68
10-25-68
10-28-68
10-29-68
10-30-68
10-30-68
11- 1-68
11- 1-68
11- 4-68
11- 7-68
11- 8-68
11- 9-68
11-11-68
11-20-68
11-26-68
12- 3-68
12-13-68
12-18-68
12-26-68
1- 2-69
1- 8-69
1-15-69
1-24-69
_
-
_
_
_
-
_
_
-
_
_
_
_
_
_
_
95.8
97.8
99.4
99.3
111.5
100.5
103.0
105.5
105.8
104.5
103;3
100.0
91.5
90.3
90.0
88.3
88.5
88.3
89.0
89.5
89.5
90.8
93.4
94.8
95.0
96.3
96.0
99.0
101.3
101.1
99.6
99.6
Flow Rate
Opening #1
2.0
2.4
2.5
3.0
3.0
3.8
3.8
3.0
2.2
2.0
1.8
2.0
2.0
2.0
2.2
2.2
2.1
2.2
2.2
2.1
2.3
2.9
2.1
2.3
2.5
2.4
2.3
2.5
(gpm)
Opening #2
6.0
5.2
5.2
4.5
4.5
4.5
5.0
5.1
3.6
5.0
3.6
1.8
2.2
2.3
2.3
2.3
2.4
2.5
2.9
2.8
2.9
3.0
3.0
3.1
3.5
3.2
3.5
3.4
Remarks
Height of drift is 84 inches.
Commenced grouting
Performed DOC Grout - Opng. #2
8:00 a.m.
2:00 p.m.
Before regrouting Opng. #1
Finished grouting
-------
TABLE 5 - Cont'd.
WATER MONITORING DATA - MINE NO. 40-016
Fluid Level - (inches) Flow Rate - (gpm)
Date Opening #1 Opening #2 Opening #1 Opening #2 Remarks
1-30-69 107.0 102.6 3.2 4.2
2- 4-69 116.3 110.3 3.3 5.0
2-10-69 117.5 111.3 3.4 4.8
2-18-69 119.0 114.3 4.3 4.8
2-24-69 116.5 111.5 3.7 4.5
3- 3-69 113.3 108.5 3.4 4.6
3-10-69 110.3 105.8 3.4 4.1
3-26-69 104.5 100.3 2.8 3.4
4- 2-69 105.0 100.8 3.1 3.6
4- 7-69 107.3 102.5 3.1 4.0
4-14-69 109.5 105.0 3.0 4.0
4-21-69 108.8 104.3 3.2 3.8
4-28-69 109.8 105.3 3.8 4.4
5- 5-69 109.3 104.8 3.6 3.9
5-12-69 112.0 107.5 3.8 5.2
5-19-69 114.3 110.0 3.4 4.4
5-26-69 112.5 108.3 3.6 5.2
6- 2-69 110.5 106.5 3.2 3.8
6-10-69 107.5 103.3 3.2 4.0
6-16-69 105.5 101.5 3.0 3.6
6-23-69 104.0 99.8 3.0 3.5
6-30-69 102.5 98.0 2.9 3.4
7- 7-69 101.0 96.5 2.8 3.4
7-21-69 101.0 96.8 2.7 3.4
7-28-69 102.8 98.5 3.2 4.4
8- 4-69 105.0 100.5 3.4 4.2
8-11-69 110.0 105.5 3.9 4.6
8-18-69 114.0 109.8
-------
TABLE 5 - Concluded
WATER MONITORING DATA - MINE NO. 40-016
Date
8-25-69
9- 2-69
9-16-69
9-29-69
10- 6-69
10-13-69
Fluid Level - (inches)
Opening #1
114.5
112.0
107.8
103.5
102.5
101.5
Opening #2
110.0
107.0
103.3
99.5
98.0
97.5
Flow Rate
(gpm
Opening #1Opening #
2.9
3.3
3.3
3.4
4.2
4.1
Remarks
Leak through highwall 25 to 30
feet right of Opening No. 1 near
base of coal seam
-------
TABLE 6
WATER ANALYSES - MINE NO. 40-016 - OPENING NO. 1
ro
-vl
Date
Sampled
9-20-68
10- 4-68
10- 7-68
10-17-68
10-21-68
10-24-68
10-25-68
10-28-68
10-29-68
10-30-68
10-31-68
11- 1-68
11- 7-68
11- 8-68
11-11-68
11-20-68
1-1-26-68
12- 3-68
12-13-68
12-18-68
12-26-68
1- 2-69
1- 8-69
1-15-69
1-24-69
1-30-69
2- 4-69
2-10-69
Cond.
(umhos/cm)
2740
2472
2650
2500
2640
2520
2560
2484
3180
2892
2944
2920
3140
3260
3100
2940
2900
3020
2844
2880
2840
2800
2784
2580
2590
2800
2680
2560
Acidity
as CaCO_
(mg/1)-3
3.2
5.2
6.1
5.4
5.5
3.7
3.4
3.7
4.8
3.6
3.7
3.9
3.1
3.1
5.8
5.2
5.4
5.6
3.6
4.5
6.0
6.0
6.1
5.9
5.0
5.8
5.8
5.9
170
310
590
400
235
130
235
200
210
200
155
220
385
300
415
290
180
130
110
120
100
30
5
20
5
110
48
123
Alka.
as CaCO,
Hardness
as
(mg/D'
0
6
41
6
20
0
0
0
3
0
0
0
0
0
49
13
51
19
0
0
82
112
73
70
11
53
80
63
1610
1510
1340
1425
1590
1640
1600
1600
1630
1675
1720
1680
1340
1640
1570
1580
1780
1690
1620
1650
1940
1730
1690
1750
1260
1670
1630
1635
Iron
(mg/1)
Sulfate
Ong/1)
149
50
192
142
143
124
144
144
144
152
166
148
163
154
178
151
110
113
101
98
82
55
67
73
64
no
72
31
1690
1508
1690
1664
1560
1690
1664
1768
1820
2028
1950
1924
1638
1768
1820
1846
1768
1742
1664
1794
1586
1664
1612
1638
1470
1480
1575
1518
Aluminum
(mg/1)
0.4
54.0
0
8.2
10.2
27.6
11.2
9.2
23.2
16.0
16.0
0
1.8
0
10.8
5.8
6.0
4.0
4.0
1.0
0
0
2.1
0
7.5
72.0
0
<2.0
-------
TABLE 6 - Concluded
WATER ANALYSES - MINE NO. 40-016 - OPENING NO. 1
PO
oo
Date
Samp! ed
2-18-69
2-24-69
3- 3-69
3-10-69
3-17-69
3-27-69
4- 2-69
4- 7-69
4-14-69
4-21-69
4-28-69
5- 5-69
5-12-69
5-19-69
5-26-69
6- 2-69
6-10-69
6-16-69
6-23-69
6-30-69
7- 7-69
7-21-69
7-28-69
8- 4-69
8-11-69
Cond.
(umhos/cm)
2400
2476
2496
2400
2400
2390
2380
2420
2314
2300
2150
2260
2240
2128
2080
2150
2146
2206
2160
2160
2178
2308
2365
2184
2190
5.9
5.9
6.5
6.4
6.7
3.6
6.0
6.1
5.1
6.0
5.6
5.8
5.9
5.6
5.9
5.7
5.3
5.6
5.9
4.2
5.8
3.8
3.3
3.8
3.5
Acidity
as CaCO,
(mq/ir
363
375
600
312
203
142
107
162
48
22
50
25
30
36
81
31
51
63
100
105
130
185
200
435
325
Alka.
as CaCO,
Hardness
as CaCO,
(mq/l)J
47
93
100
109
52
58
60
83
9
71
37
55
58
0
45
30
67
65
40
0
54
0
0
0
0
1650
1000
1185
1395
1550
1235
1350
1360
1405
1425
1755
1360
1600
1250
1290
955
1330
1295
1560
27
1360
1360
305
1355
1430
Iron
(mg/1)
55
55
108
73
99
69
81
79
90
66
90
76
68
74
61
79
9
80
96
113
96
98
152
95
132
Sulfate
(mg/1)
1500
1320
1505
1575
1400
1470
1400
1575
1400
1295
1365
1400
2030
1350
1400
1425
1425
1450
1550
1250
1350
1260
1290
1425
1600
Al urni num
(mg/1)
0.7
1.0
<2.0
<2.0
<4.0
<5.0
<5.0
7.0
0.6
1.4
0.4
0.7
2.4
0
1.7
1.1
1.2
0.6
0.1
0
0
.5
.6
.0
4.0
-------
TABLE 7
WATER ANALYSES - MINE NO. 40-016 - OPENING NO. 2
10
Date
Sampled
9-20-68
10- 4-68
10- 7-68
10-17-68
10-21-68
10-24-68
10-25-68
10-28-68
10-29-68
10-30-68
10-31-68
11- 1-68
11- 7-68
11- 8-68
11-11-68
11-20-68
11-26-68
12- 3-68
12-13-68
12-18-68
12-26-68
1- 2-69
1- 8-69
1-15-69
1-24-69
1-30-69
2- 4-69
2-10-69
Cond.
(umhos/cm)
2518
2426
2564
2436
2574
2680
2460
2420
3140
2892
2844
2860
2640
2660
2822
2840
2860
2860
2684
2820
2660
2600
2680
2440
2500
2700
2600
2280
pH
4.7
5.7
6.2
6.0
5.8
3.1
4.1
4.9
5.8
5.0
4.8
6.0
5.8
5.8
6.3
6.7
6.3
6.4
6.0
6.0
6.5
6.5
6.3
6.1
6.5
6.2
6.7
6.6
Acidity
as CaCO.,
(mg/1)3
170
260
630
280
140
190
115
180
90
125
140
90
15
30
25
5
120
0
30
10
10
20
0
10
410
70
101
238
Alka.
as CaC03
(mg/1)
0
27
57
36
33
0
0
17
40
0
0
32
41
42
14
37
148
131
158
125
201
204
152
128
190
102
173
191
Hardness
as CaCOo
(mg/ir
1570
1520
1250
1490
1500
1575
1630
1540
1670
1565
1655
1565
1660
1700
1710
1810
1700
1800
1650
1900
1730
1790
1800
1660
1100
1710
1430
1220
Iron
(mg/D
131
163
136
116
110
143
108
120
127
142
139
120
98
96
91
82
78
76
58
78
49
44
60
67
39
66
54
48
Sulfate
(mg/1)
1560
1534
1612
1560
1638
1560
1404
1664
1690
1898
1924
1885
1404
1690
1716
1820
1560
1716
1560
1664
1508
1560
1508
1560
1560
1470
1540
1608
Al umi num
(mg/1)
9.2
10.0
2.2
6.6
2.4
20.2
17.0
11.6
18.0
0
18.0
0
7.0
8.0
2.4
1.8
4.4
4.4
7.8
7.4
0
0
3.8
0
<2.0
<2.0
0
<2.0
-------
TABLE 7 - Concluded
WATER ANALYSES - MINE NO. 40-016 - OPENING NO. 2
CO
o
Date
Samp! ed
2-18-69
2-24-69
3- 3-69
3-10-69
3-17-69
3-27-69
4- 2-69
4- 7-69
4-14-69
4-21-69
4-28-69
5- 5-69
5-12-69
5-19-69
5-26-69
6- 2-69
6-10-69
6-16-69
6-23-69
6-30-69
7- 7-69
7-21-69
7-28-69
8- 4-69
8-11-69
Cond.
(umhos/ctn)
2320
2334
2340
2240
2260
2240
2332
2440
2364
2280
2220
2214
2250
2120
2008
2110
2132
2112
2158
2038
2112
2224
2176
2128
2244
pH
6.7
6.4
6.9
6.5
6.7
6.2
6.4
6.3
6.7
6.7
6.5
6.4
6.5
6.2
6.4
6.2
5.9
6.0
6.0
6.3
6.1
5.9
6.3
6.3
6.5
Acidity
as CaC03
112
75
425
262
159
30
17
9
0
155
no
21
56
10
93
27
89
0
27
65
75
0
17
875
163
Alka.
as CaCO,
(mg/l)J
181
160
183
123
100
127
121
160
140
185
171
155
156
171
153
109
127
118
97
81
112
84
88
117
177
Hardness
as CaCO,
(mg/ir
1205
920
1125
1265
1275
1445
1405
1530
1570
1150
1835
1355
1320
1195
1100
915
1245
1325
1430
1235
1355
1675
1520
1575
1490
Iron
(mg/1)
31
49
81
50
73
60
57
52
51
33
29
43
43
28
45
55
68
63
66
72
59
70
64
62
52
Sulfate
(mg/1)
1380
1380
1295
1365
1435
1400
1505
1505
1400
1260
1120
1260
1890
1275
1500
1375
1350
1250
1400
1115
1225
1260
1250
1200
1575
Al um'i num
(mg/1 )
0.5
1.0
<2.0
<2.0
<8.0
<5.0
<5.0
<5.0
0.4
1.3
0.3
0.5
2.1
0
1.1
1.2
0.8
0.5
0.2
0
0
0
0.1
0.2
0.6
-------
TABLE 8
PNEUMATIC CONVEYING TEST
3/4-INCH AGGREGATE
Test
No.
1
2
3
4
5
6
7
8
9
10
11
12
13
14
15
16
17
18
19
20
21
22
23
24
25
26
27
28
29
30
Aggregate
Conveyed
(Ibs)
4820
4780
4605
4550
4270
4310
4220
4220
4200
4200
4180
4200
4180
4150
4130
4150
4120
4080
4000
4000
4000
4000
3980
6070
6380
6370
6290
6180
6120
6100
Conveying
Time
(min)
35.0
29.0
36.0
19.0
15.0
7.0
13.0
10.0
8.0
6.0
7.5
4.0
7.0
6.0
5.0
4.0
4.5
4.0
4.1
3.7
5.0
4.0
5.5
21.5
21.0
17.0
11.0
10.5
9.4
9.3
Conveying
Rate
(Ibs/min)
138
165
128
239
285
616
325
422
525
700
557
1050
597
692
826
1038
916
1020
976
1081
800
1000
724
282
304
375
572
589
651
660
Initial Tank
Pressure
(psig)
15
15
15
15
15
15
15
15
15
15
15
15
15
15
15
15
15
15
15
15
15
15
15
15
15
15
20
20
20
20
Conveying
Line
(ft)
80
80
80
80
80
80
80
80
80
80
40
40
40
40
40
40
60
60
60
80
80
80
80
80
80
80
80
80
80
80
Note: Air supply for this test was 460 scfm (±5%).
was a 4-inch ID rubber hose.
The conveying line
-131-
-------
TABLE 9
Properties:
CJ
ro
Composition:
Cost - Material:
Cost - Equipment:
ORGANIC FOAM MATERIALS INVESTIGATION
Foam A1 c~m °2
Foam BJ
Bulk Density - 2 to 2.5 pounds per
cubic foot minimum (3.0 pounds
per cubic foot minimum possible
if ambient temperature is no
lower than 60°F)
Compressive Strength at Yield
Point - 48 psi
Tensile Strength - 70 psi
Resistance to Dilute Acids - Good
Resistance to Dilute Alkali - Exc.
Resistance to Degradation by
Molds & Fungi - Good
Polyisocyanate, preheated to
130-150°F
Polyol, containing flourocarbon
and surfactant - 72°F
$2.40 per cubic foot - spray
applied at 3.0 pounds per
cubic foot
$2,383.00 for Portable Spray
Application
Bulk Density - 1/4 to 2 pounds per
cubic foot
Compressive Strength - 3 to 5 psi
Water Absorption - Very high
Resistance to Dilute Acids - Poor
Resistance to Biological Attack - Poor
Resistance to Water - Poor
Urea-formaldehyde tesin - partly
polymerized
Monobasic ammonium phosphate - catalyst
Casein - stabilizer
$2,500.00 - $3,000.00 for Foam
Generating Unit
1 Polyurethane
2 Urea-formaldehyde
-------
TABLE 10
INORGANIC FOAM TESTS
00
CO
Components1
SOLUTION A3
Sodium Silicate - Grade 40
Formaldehyde Solution, 37%
Water
Howco Suds'*
Silica Flour
Fly Ash
SOLUTION B3
Sodium Silicate - Grade 40
Hydrogen Peroxide Solution, 30%
Water
Howco Suds'*
Silica Flour
Fly Ash
Final Foam Volume as % of
original fluid volume
Foam Condition5
Formula No.
1360-53-1
333.5
102.0
564.5
333.5
69.0
564.5
230%
Fragile -
crumbles
easily.
Formula No.
1360-53-2
333.5
102.0
564.5
1.0
333.5
69.0
564.5
1.0
350%
Very
fragile -
crumbles
very easily.
Formula No.
1360-55-1
50.0
15.3
50.0
10.4
5.0
153%
Much
stronger
than
1360-53-1
Formula No.
1360-62-12
62.5
25.0
62.5
250 g
66.3
17.5
66.3
250 g
280%
Good
strength -
permeable
to water.
Components shown as parts by volume except as noted.
100 cc of each solution mixed.
Equal volumes of Solution A and Solution B mixed in each formula.
Halliburton trade name for a water soluble biodegradable surfactant blend.
Hydrogen gas liberated during generation of foam on all tests.
Formula No.
1360-64-32
62.5
25.0
62.5
250 g
66.3
17.5
66.3
250 g
203%
Good
strength,
-------
I
CJ
I
Slurry
TABLE 11
INITIAL TEST DATA - SLURRY COMPOSITION
Components
Remarks
1 100 parts portland cement1, 8 parts gypsum, Rejected. Not set in 1 hour - medium angle of
12 parts bentonite and 109 parts water repose.
2 100 parts cement, 12 parts bentonite and
79 parts water
3 100 parts cement, 8 parts gypsum and 46
parts water
4 100 parts cement, 8 parts gypsum, 3 parts
calcium chloride and 46 parts water
5 20 parts cement, 40 parts fly ash, 40 parts
gypsum and 41 parts water
6 75 parts cement, 25 parts gypsum and 38
parts water
7 100 parts cement, 10 parts gypsum, 10 parts
bentonite and 3 parts calcium chloride pre-
mixed in 92 parts water
8 20 parts cement, 40 parts fly ash, 40 parts Selected. Set in 10 minutes. Medium angle of
gypsum, 4 parts bentonite and 60 parts water repose.
Rejected. Not set in 3 hours - low angle of
repose.
Rejected. Not set in 1 hour - low angle of
repose.
Selected for further consideration. Set in 1
hour - medium angle of repose.
Rejected. Set in 15 minutes. Low angle of
repose.
Selected. Set in 10 minutes. Medium angle of
repose.
Rejected. Angle of repose too low.
1 All cement used in slurry tests was ASTM Type I
-------
TABLE 11 - Continued
INITIAL TEST DATA - SLURRY COMPOSITION
CO
on
SIur _ry_ Components
9 Same dry materials as Slurry No. 8. Ben-
tonite was prehydrated in 40 parts water.
Other materials were mixed with 40 parts
water and two slurries blended
10 Same dry materials as Slurry No. 8. Ben-
tonite and gypsum mixed in 50 parts water.
Other materials mixed in 30 parts water
and two slurries blended
11 Same as Slurry No. 10 except cement and fly
ash mixed in 20 parts water
12 Same as Slurry No. 10 except bentonite and
gypsum mixed in 46 parts water and other
materials mixed in 18 parts water before
blending slurries
13 20 parts cement and 40 parts fly ash mixed
in 40 parts water. 40 parts gypsum and 4
parts bentonite mixed in 26 parts diesel
oil, then two slurries blended
14 40 parts cement, 40 parts fly ash, 40 parts
gypsum, 4 parts bentonite and 60 parts water.
Bentonite prehydrated in 60 parts water
Remarks
Selected. Set in less than 30 minutes
medium angle of repose.
Rejected. Low angle of repose.
Rejected. Low angle of repose.
Selected. Set in 30 minutes - medium angle
of repose.
Selected. Very viscous in less than 1 minute.
Rejected. Faulty sample. Fly ash would not
blend into prehydrated gel slurry
-------
TABLE 11 - Continued
INITIAL TEST DATA - SLURRY COMPOSITION
CO
O)
Slurry Components
15 20 parts cement, 40 parts fly ash, 40 parts
gypsum, 2 parts bentonite and 86.5 parts
water. Gypsum mixed in 24 parts water and
other dry materials mixed in 62.5 parts
water, then slurries blended
16 20 parts cement, 40 parts fly ash, 40 parts
gypsum and 4 parts bentonite. Bentonite pre-
hydrated in 80 parts water. Cement and fly
ash added, then gypsum
17 Same as Slurry No. 8 except that cement, fly
ash and water were mixed together, then
blended gypsum and bentonite added quickly
to slurry
18 20 parts cement, 40 parts fly ash, 40 parts
gypsum, 2 parts bentonite and 80 parts water.
Bentonite prehydrated in 30 parts water, then
other materials added separately into slurry
19 20 parts cement, 40 parts fly ash, 40 parts
gypsum, 2 parts bentonite, 2.4 parts salt and
80 parts water. Bentonite prehydrated in 30
parts water. Other dry materials mixed in 50
parts water and two slurries blended.
20 100 parts cement, 10 parts gypsum, 12 parts
bentonite and 110 parts water
Remarks
Rejected. Too thin with low angle of repose.
Rejected. Did not obtain good distribution
of gypsum.
Rejected. Became too viscous to handle before
thoroughly mixed.
Rejected. 3 hours setting time.
Rejected. Set too slowly.
Rejected. Not set in 1 hour.
-------
TABLE 11 - Concluded
INITIAL TEST DATA - SLURRY COMPOSITION
CO
««J
I
SIurry Components
21 100 parts cement, 10 parts gypsum, 6 parts
bentonite, 3 parts calcium chloride and
110 parts water. Bentonite mixed in 60
parts water. Other dry materials mixed in
50 parts water and alurries blended
22 60 parts cement, 40 parts fly ash, 6 parts
bentonite and 100 parts water. Bentonite
mixed in 55 parts water. Other dry mate-
rials mixed in 45 parts water and slurries
blended
23 60 parts cement, 40 parts fly ash, 6 parts
bentonite and 110 parts water. Bentonite
mixed in 60 parts water. Other dry mate-
rials mixed with 50 parts water and
slurries blended
24 60 parts cement, 40 parts fly ash, 6 parts
bentonite, 4 parts calcium chloride and
110 parts water. Mixed same as Slurry No.
23 except that calcium chloride was added
to 50 parts water before mixing dry mate-
rials or blending slurries
Remarks
Selected. Very viscous but not quick-setting,
Selected.
viscous.
Set in less than 1 hour and very
Rejected. Not sufficiently viscous.
Rejected. Not sufficiently viscous.
-------
TABLE 12
ADDITIONAL TEST DATA - SELECTED SLURRIES
SI urry
4
6
8
9
12
13
21
22
30 min
Firm
Firm
160
56
1502
Firm
Firm
Firm
Strength,
1 hr 2 hrs
Firm 217
206 240
-
46
-
105
-
18 35
psi , at Various Time Intervals1
3 hrs
400
-
140
30
145
-
12
_
5 hrs
925
-
185
38
138
-
50
_
24 hrs
3525
1965
315
100
275
170
300
261
48 hrs
4675
2840
-
130
-
-
660
380
72 hrs
-
3000
-
170
-
-
725
445
Material.
Cost/cu yd
$ 31.10
32.43
16.80
14.00
15.20
24.35
18.28
18.00
1 Compressive strength specimens were placed in a water bath at room temperature until tested. Test
was conducted on 2-inch standard test cubes in accordance with Bulletin API RP 10B of the American
Petroleum Institute.
2 Strength shown is for 40 minutes' time.
-------
TABLE 13
QUICK-SETTING SLURRY COMPOSITIONS
Formula No.
dvitiuMian+e1 1360-61-1
«J
CO
<£>
1
wvi'XyL™mgj*. _n
CEMENT SLURRY
Portland Cement
Flv Ash
r i JF f^jM
Silica Flour
Hater
Sand
Howe o- Suds2
Slurry Height, Ib/gal
SILICATE SLURRY
Sodium Silicate1
Silica Flour
Bentonlte*
Sand
JQIIU
Hater
Slurry Weight, Ib/gal
Combined Slurry Ut.
Ib/gal
Approximate Setting
Time, seconds
Estimated Material Cost
per cubic yard
71.43
.
28.57
16.23
52.63
•*
47.37
19.82
13.02
20
$23.77
Formula No.
1360-67-1
35.71
35.71
28.58
-
15.29
61.94
—
38.06
9.8
12.54
29
$17.56
Formula No.
1360-69-1
32.62
23.14
44.24
-
13.16
33.82
35.80
•
30.38
12.24
12.70
35
$18.10
Formula No
1360-72-1
21.74
-
15.42
29.49
33.35
15.21
6.55
27.74
2.91
38.78
24.02
15.17
15.19
28
$13.10
Formula No.
1414-11-1
26.05
26.05
17.30
30.60
-
14.87
60.0
3.0
•
37.0
10.29
12.58
23
$17.94
Formula No.
1414-13-1
21.74
-
15.42
29.49
33.35
15.21
7.89
3.51
««•«
.71
41.89
12.51
13.86
24
$ 9.47
Formula No.
1414-17-1
66.67
—
33.33
—
15.27
18.69
62.23
0.39
•
18.69
15.00
15.14
8
$26.97
Formula No.
1414-18-1
66.67
~
33.33
"
15.27
22.75
52.77
1.73
~
22.75
14.08
14.68
8
$26.26
Formula No
1414-26-1
66.34
"
33.16
Ocn
• SHJ
15.27
22.75
52.77
1.73
~
22.75
14.08
14.68
8
$32.68
1 Components shown as per cent by weight
2 Halliburton trade name for a water soluble biodegradable surfactant blend
1 Grade 40
* Bentonlte prehydrated In water prior to adding other components
-------
TABLE 14
DATA FOR SELECTED SILICATE CEMENT SLURRY
Material Per Yard of Final Slurry1
Slurry2
1
1
2
2
2
Material Sacks
Portland Cement 12
Water
Bentonite
Water
Sodium Silicate
Grade 40
Gallons
-
58
-
75
25
Pounds
1128
483
50
625
292
1 Cost of final slurry for materials only calculated to be $20.90.
2 Slurry 1 and Slurry 2 mixed in equal volumes to obtain final slurry.
Compressive Strength - Final Slurry
2 Hours 18 Hours 24 Hours 48 Hours 72 Hours
100 psi 535 psi 990 psi 555 psi3 925 psi3
3 Discrepancy thought to be because of difficulty in blending slurries
and obtaining sample due to fast set.
-140-
-------
TABLE 15
tOW TEMPERATURE TESTS OF BENTONITE-SODIUM SILICATE SLURRY
Sodium
Test Mix Ratio of Bentonlte Bentonlte Bicarbonate
Temperature1 Temperature* to Sodium Silicate (i) (%) Remarks
30° F M°F 3*>' 8 o Little viscosity increase -
not set In one hour
30° F X' f 3 f 1 8 i.o Little viscosity Increase -
not set In one hour
40° F 45° F 3 to 1 8 l.o Thin - not set In 1 minute -
self-supporting* In 5 Minutes
400 F 45" F 2 to 1 8 0 Viscous In 25 seconds - self-
supporting In 1 Minute
40° F 45° F 3 to 1 10 0 Not setting 1n 1 minute - self-
supporting In 3 Minutes
40° F 45* F 3 to 1 10 1.0 Not setting In 1 Minute - self-
supporting In 3 Minutes.Thinner
than slurry above.
40° F 45° f 2 to 1 10 0 Viscous Immediately - Initial
set" In 20 seconds
40° F «° F 2 to 1 10 1.0 Thin for 15 seconds - Set in
20 seconds
40° F 45' F 3 to 1 12 0 Viscous - not self-supporting
In 2 Minutes
«• F 45' F 3 to 1 12 1.0 Thin for 20 seconds - set In
30 seconds
«0° F 45° F 2 to 1 12 0 Viscous - Initial set In 30
seconds
50° F 54° F 3 to 1 8 0 Viscous but not set In 10
Minutes
50° F 54" F 3 to 1 8 1.0 Thin - Initial set In 30
seconds
50° F 54° F 2 to 1 8 0 Viscous - Initial set In 30
seconds
50° F 54" F 3 to 1 10 0 Viscous - self-supporting In
2 nlnutes
SO0 F 54s F 3 to 1 10 1.0 Thin for 15 seconds - Initial
set 1n 20 seconds
SO" F 54° F 2 to 1 10 0 Viscous - Initial set In 20
seconds
1 The bentonlte for these tests was mixed In 50° F water at a slow speed on a Olspersator mixer for 20
minutes and then the sodium silicate was added. This slurry and the cement slurry were then cooled
to the test temperature Indicated before being mixed together at a volumetric ratio of 1 to 1.
* Temperature after slurries were mixed together.
1 Estimated time, by appearance and feel of the material, from time of nixing until It would support
Itself to permit continued building of a bulkhead.
* Time from Initial mixing of the 2 slurries until no more stirring could be accomplished with a
spatula.
-141-
-------
TABLE 16
WATER MONITORING DATA - MINE NO. 62-008 SPECIFIED OPENINGS
-P.
ro
t
Date
9-20-68
9-21-68
10- 1-68
10-17-68
10-29-68
11- 8-68
11-20-68
11-22-68
11-26-68
12- 3-68
2-12-69
2-13-69
Flow
(gpm)
30
-
48
68
60
60
156
-
98
122
_
-
Flow
(gpm)
_
2.8
-
_
_
-
_
0
-
0
_
-
Head
(in.)
_
*
-
_
-
-
_
**
-
-
_
-
Flow
(gpm)
.
-
-
_
_
-
_
-
-
-
15
-
Head
(in.)
.
*
-
_
_
-
_
-
-
-
_
t
Flow
(gpm)
_
-
-
_
_
-
_
_
_
_
_
Remarks
* No bulkheads built yet in any opening.
** Built silicate cement bulkhead 8 feet
back in opening.
t Rear silicate cement bulkhead built 28
2-15-69
2-27-69
2-28-69
1 Sample Point #1
2 Sample Point #4
3 Sample Point #5
" Sample Point #8
tt
21.0
feet from entrance.
tt Front silicate cement bulkhead built.
Grout pipes installed into void between
bulkheads
Grouted void between bulkheads-S.P. #5.
Head increased from 10.5 inches to 21
inches in 20 hours - opened drain line.
outlet of 48-inch culvert at tipple - Mine No. 62-008.
main Pittsburgh portal designated as Opening No. 4.
utility portal to right of main Pittsburgh portal - Opening No. 5.
fanway opening at right end of other openings - partially collapsed.
-------
TABLE 16 - Continued
WATER MONITORING DATA - MINE NO. 62-008 SPECIFIED OPENINGS
Date
3-13-69
3-17-69
3-26-69
4- 7-69
4-21-69
4-28-69
5- 5-69
5-12-69
5-19-69
5-26-69
6- 2-69
6-16-69
6-23-69
6-30-69
7- 7-69
7-21-69
7-28-69
8- 4-69
8-11-69
8-18-69
C D dtfl
o • r • TT i
Flow
(gpm)
1 3r '
—
_
_
M
—
H
_
_
_
_
_
_
_
_
_
_
_
_
-
1 S.P.
Flow
(gpm)
__*__Ylf f .. _.r*
0
0
0
0
0
0
0
0
0
0
0
0
0
0
0
_
0
0
0
Seep
#42
Head
(in.)
_
-
-
c -
42.5
42.0
43.3
48.3
49.0
48.8
48.8
48.8
48.5
48.5
49.5
50.0
50.8
51.8
53.0
S.P.
Flow
(gpm)
10.2
0
0
0
0
0
0
0
0
0
0
0
0
0
0
_
0
0
0
0
#53
Head
(in.)
28.5
24.0
32.5
35.0
38.5
39.0
41.0
43.8
44.5
44.8
44.3
44.8
44.5
44.5
45.8
46.5
47.0
48.0
49.8
S.P. #8*
Flow
(qpm) Remarks
Closed drain lines on both bulkheads.
Flowing from S.P. #8 but no measurement taken.
S.P. #8 flowing but not measured.
-
-
8.6
7.3
12.5
10.9
8.7
7.5
3.0
2.3
1.8
1.8
2.3
2.7
2.6
3.1
3.0
Sample Point #1 - outlet of 48-inch culvert at tipple - Mine No. 62-008.
Sample Point #4 - main Pittsburgh portal designated as Opening No. 4.
Sample Point #5 - Utility portal to right of main Pittsburgh portal - Opening No. 5.
Sample Point #8 - fanway opening at right end of other openings - partially collapsed.
-------
•PI
-£»
I
TABLE 16 - Concluded
WATER MONITORING DATA - MINE NO. 62-008 SPECIFIED OPENINGS
S.P. aPI1 S.P. #42 S.P. #53 S.P. #B*
FlowFlowHead FlowHead Flow
Date (gpm) (gpm) (In.) (gpm) (in.) (gpm) Remarks
8-25-69 - - 48.0 0 50.5 3.0
9- 2-69 - - 54.5 - 50.8 3.2
9-16-69 - - 54.5 - 51.0 1.9
9-22-69 - - 54.3 - 50.8 2.0
9-29-69 - - 54.0 - 50.3 1.7
10- 6-69 - - 54.0 - 50.3 1.2
10-13-69 - - 54.3 - 50.3 1.0
Sample Point #1 - outlet of 48-inch culvert at tipple - Mine No. 62-008.
Sample Point #4 - main Pittsburgh portal designated as Opening No. 4.
Sample Point #5 - utility portal to right of main Pittsburgh portal - Opening No. 5.
Sample Point #8 - fanway opening at right end of other openings - partially collapsed,
-------
TABLE 17
tn
Date
Sampl ed
9-20-68
10- 1-68
10-17-68
10-29-68
11- 8-68
11 -20-68
11-26-68
12- 3-68
9-21-68
12-13-68
Location
S.
S.
S.
S.
S.
S.
S.
S.
S.
S.
P.
P.
P.
P.
P.
P.
P.
P.
P.
P.
No.
No.
No.
No.
No.
No.
No.
No.
No.
No.
Cond.
(umhos/cny) 2tt
I1
1
1
1
1
1
1
1
42
53
2854
2690
2620
2560
2350
2120
2444
2426
5210
3060
3
4
3
4
3
3
3
3
3
2
.5
.5
.6
.3
.4
.9
.9
.9
.0
.8
WATER ANALYSES - MINE NO. 62-008
Acidity Alka. Hardness
as CaCOo as CaCOo as CaCXh
(mg/1) (mg/1) (mg/1)
180
350
365
190
260
350
280
250
1170
2260
0
0
0
0
0
0
0
0
0
0
1440
1285
1180
1330
1050
905
1120
1125
1855
1235
Iron
(mg/1)
31.7
18.0
19.2
17.6
26.4
24.6
37.2
20.4
246.0
600.0
Sulfate
(mg/1)
1560
1560
1560
1560
1352
1443
1404
1378
3016
4160
Aluminum
(mg/1)
9.2
8.6
11.0
7.7
14.0
14.0
21.2
7.6
64.0
88.0
1 Sample Point No. 1 - outlet of 48-inch culvert at tipple - Mine No. 62-008.
2 Sample Point No. 4 - main Pittsburgh portal designated as Opening No. 4.
3 Sample Point No. 5 --utility portal to right of main Pittsburgh portal.
-------
TABLE 18
SCREEN ANALYSES - GREEK LIMESTONE SAMPLES
U.S. Standard
Sieve Mesh Size
Passed
1/2"2
3/8" 2
4
8
16
SO
100
Retained
1/2" 2
3/8" 2
4
8
16
50
100
Pan
Greer 11 2>
Per Cent
Retained
0
4.3
73.2
21.6
-
-
-
0.9
Greer I12B1
Per Cent
Retai ned
-
0
38.8
42.0
9.8
5.8
-
3.6
Greer 11 3 '
Per Cent
Retained
0
8.7
-
-
-
71.7
19.6
SCREEN ANALYSES - HARROLO LIMESTONE SAMPLES
U.S. Standard
Sieve Mesh Size
Passed
3/4"*
1/2"2
1/2"2
3/8" 2
3-1/2
4
8
20
100
Retained
3/4" 2
1/2"2
3/8" a
3-1/2
4
8
8
20
100
Pan
Harrold I7»
Per Cent
Retained
0
85.85
11.83
-
1.733
-
0.083
0.025
0.275
0.133
Harrold IIP
Per Cent
Retained
0
69.90
22.96
-
5.66
-
0.258
0.133
0.408
0.600
Harrold 112*
Per Cent
Retained
-
0
-
53.0
-
28.0
-
20.0
0.1
0.6
Harrold 113s
Per Cent
Retained
-
-
-
-
4.67
-
6.05
37.92
34.18
16.45
1 Greer Limestone Company, Greer, West Virginia.
2 Diameter of screen opening.
Paul Harrold Limestone Quarry, Holf Summit (Clarksburg, U. Ya.).
-146-
-------
TABLE 19
RETENTION TIME FLOW TEST - HARROLD #12 LIMESTONE1
Sample
No.
1
2
3
4
5
6
7
8
Total
Time
(hire)
1
8
24
48
72
124
158
209
Retention
Time
(mins)
23.86
26.90
23.86
17.30
21.20
66.10
30.00
27.50
£H
7.50
7.40
7.45
7.20
7.30
7.60
7.40
6.60
Alkalinity
as CaCOo
(mg/ir
180
160
119
80
100
119
80
39
Total
Iron
(mg/1)
10
10
10
5
5
5
40
—
Dissolved
Iron
(mg/1)
<1.7
<0.5
<0.5
<0.5
<0.5
<0.5
<0.5
<0.5
Ca++ SO*
475
575
575
575
575
600
575
575
(mg/1)
1450
1450
1450
1450
1400
1560
1400
1450
1 Acid mine water used in test had pH of 3.2, acidity of 180 mg/1 as CaC03> total iron of 30 mg/1
with 500 mg/1 calcium and 1450 mg/1 sulfate.
-------
TABLE 20
FLOW TEST DATA - MODIFIED GREER #13 GRADED LIMESTONE1
Total Alkalinity Total Dissolved
Sample Time as CaCO^ Iron Iron Ca++ SO/
No.
1
2
3
4
(hrs)
1
24
100
125
fill .
7.96
7.90
7.86
7.60
(mg/ir
160
239
200
160
(rag/1)
20
60
10
20
(mg/1)
<0.5
<0.5
<0.5
<0.5
500
626
625
626
(mg/1) '
1550
1450
1450
1400
1 Acid mine water used in test had pH of 3.2, acidity of 180 mg/1 as
» total iron of 30 mg/1 with 500 mg/1 calcium and 1450 mg/1
sulfate.
TABLE 21
FLOW TEST DATA - MODIFIED GREER #13 LIMESTONE CONTAINING 1% BaCO^
Total Alkalinity Total Dissolved
Sample Time as CaCOo Iron Iron Ca++ Spjj
No. (hrs) pH (mg/ir (mg/1) (mg/1) (mg/1)
1 1 8.00 239 10 <0.5 625 1450
2 24 7.85 219 10 <0.5 625 1450
3 100 8.00 219 10 <0.5 675 1400
4 125 8.20 259 10 <0.5 675 1400
i Acid mine water used in test had pH of 3.2, acidity of 180 mg/1 as
CaC03, total iron of 30 mg/1 with 500 mg/1 calcium and 1450 mg/1
sulfate.
-148-
-------
TABLE 22
FLUID FLOW DATA
PERMEABLE PLUG TEST NO. 1
Flow Measurement (gpm)
Standplpe Fluid Levels (Inches)
vo
Date
ZwZS.
12-20-68
12-23-68
12-24-68
12-27-68
12-30-68
12-31-68
1- 2-69
1- 3-69
1- 6-69
1- 7-69
1- 8-69
1- 9-69
1-10-69
1-13-69
1-13-69
1-14-69
1-15-69
1-16-69
1-17-69
1-20-69
1-21-69
1-24-69
1-27-69
1-28-69
1-29-69
1-30-69
1-31-69
2- 3-69
2- 4-69
2- 5-69
2- 6-69
Total
Mine Flow
28.2
.
28.2
27.9
.
.
_
_
.
.
28.2
-
.
m.
_
_
„
•
.
•
28*2
.
—
.
.
.
.
.
-
Compartment No.l
Overflow
.
.8
13.5
16.2
15.0
20.5
24.4
22.6
26.0
27.8
27.1
27.0
27.0
24.0
21.9
28.2
24.0
24.1
26.8
26.3
27.6
27.6
27.1
31.9
32.0
48.7
42.4
39.6
-
Input
9.4
,
.
.
-
1.2
_
3.3
3.3
.7
.7
.4
.6
_
.3
.3
.2
.2
.2
0
.3
0
0
.3
.3
.1
0
.2
.2
.2
-
Discharge
7.5
8.5
3.0
3.0
3.8
2.3
.1
.9
.3
.2
.1
.3
.3
_
.3
.2
.2
.2
.2
.3
.2
.1
.2
.2
.2
.1
.1
.2
.5
-
Compartment No. 2
Input
9.4
.
_
.
-
3.0
-
1.6
1.6
.4
.4
.2
.1
_
,3
.3
.1
.2
.1
0
.3
.1
0
.3
.3
.1
.1
0
.2
.2
-
Discharge
6.0
7.5
5.0
2.5
2.3
1.8
.3
.8
.4
.2
.3
.2
.1
.
.2
.3
.2
.3
.3
.5
.2
.1
.3
.3
.1
.1
.1
.1
.6
-
Compartment No. 3
Input
9.4
.
.
.
-
4.2
.
3.8
3.8
1.0
1.1
.5
.5
_
3.6
5.7
5.7
3.8
3.9
1.4
1.4
.5
.6
.6
.6
1.0
.8
.6
.6
1.0
-
Discharge
5.0
5.0
6.4
3.8
5.0
2.5
1.0
.8
.7
.6
.3
.6
.4
.
3.0
3.7
3.5
2.6
.9
.8
.3
.1
.1
.1
.1
.1
.1
.1
.5
-
Comp.
1-1
39
41
57
59
60
60
-
60
41.5
41.5
36.3
53.5
57.5
.
53.5
46.0
28.5
30.5
12.0
56.0
32.5
51.0
57.5
57.0
54.5
40.5
40.0
57.0
56.5
No. 1
1-2
12.0
11.0
12.0
9.5
5.0
12.0
-
9.5
5.0
3.5
4.0
4.0
3.0
-
3.0
3.0
3.0
3.0
3.0
4.0
2.0
1.5
3.5
4.5
3.5
3.0
1.0
3.0
7.0
Comp. No. 2
2-1 Z-2
47 12.3
41 8.0
57 15.0
58 10.5
60 10.0
60 10.5
-
60 7.0
52 4.0
54 2.0
53.3 1.5
51.5 3.0
63.0 2.0
-
50.5 2.0
35.0 4.0
18.0 3.5
43.0 4.0
8.0 5.0
52.5 6.0
45.5 3.0
7.0 1.5
54.5 5.5
55.0 4.0
52.0 4.0
51.0 3.0
14.0 .5
57.0 3.5
55.5 5.0
Comp.
3-1
14
14
36
47
52
56.5
55.0
51.5
48.0
42.5
35.0
30.5
10.5
-
49.5
43.5
34.0
30.5
19.0
36.0
17.5
6.0
24.0
16.0
8.0
.6.0
4.0
4.5
55.5
No. 3
3-2
10.5
10.0
18.0
22.5
24.0
27.0
23.3
22.5
19.0
13.0
8.5
6.0
5.0
-
23.0
15.0
13.0
14.0
16.0
16.0
9.0
4.5
6.0
9.0
5.5
5.0
2.0
3.0
18.0
Comments
Set valves to constant flow
Probed box with metal rod
12 Compartment overflowed
II Compartment overflowed
Modified #1 Input line
Readjusted valves
Readjusted valves
II Compartment flooded
Cleaned and reset valves
Cleaned and reset valves
Reset valves
Locally heavy rains
Reset II valve
Locally heavy rains
Terminated test
-------
tn
O
TABLE 23
WATER SAMPLE ANALYSES - PERMEABLE PLUG TEST NO. 1 - MINE NO. 40-085
Date
Sampled
11- 6-68
11-26-68
12-12-68
12-20-68
12-20-68
12-20-68
12-24-68
12-24-68
12-24-68
12-24-68
1- 2-69
1- 2-69
1- 2-69
1- 2-69
1- 8-69
1- 8-69
1- 8-69
1- 8-69
1-15-69
1-15-69
1-15-69
1-15-69
1-24-69
1-24-69
1-24-69
1-24-69
Location
Mine Flow
Mine Flow
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Cond.
(umhos/cm)
3080
3080
2980
2840
2804
2840
3040
2680
3080
2820
2940
3100
3200
3200
3120
3240
3360
3280
2960
3060
3120
2980
3200
3200
3300
3200
M
3.3
3.7
2.9
5.9
6.1
6.2
3.8
5.8
5.9
6.0
3.8
6.1
6.2
6.2
3.2
6.3
6.3
6.2
3.3
6.0
6.2
5.7
3.2
6.6
6.6
6.6
Acidity
as CaCOs
(mg/1)
515
580
600
200
110
70
580
200
160
0
560
80
0
20
475
10
0
0
640
60
0
200
680
203
189
168
Alka.
as CaCOs
(mg/1)
0
0
0
44
67
85
0
59
68
107
0
133
182
164
0
128
159
142
0
134
177
67
0
160
220
141
Hardness
as CaCOs
(mg/1)
1165
1150
1190
1780
1390
1730
1350
1470
1590
2270
1320
1860
1880
1990
1300
1800
1870
1750
1200
1790
1670
1530
1030
2380
1520
1120
Iron
(mg/1)
124.8
122.4
120.0
110.4
120.0
60.0
156.0
108.0
120.0
64.8
129.6
117.6
66.0
84.0
121.2
62.4
48.0
62.9
129.6
64.8
40.8
98.4
25.9
38.4
21.6
26.4
Sulfate
(mg/1)
1768
1820
1820
1560
1690
1430
1820
1820
1846
1820
1820
1820
1820
1768
1846
1820
1885
1934
1846
1885
1820
1794
1840
2000
1845
2000
Aluminum
(mg/1)
25.6
19.6
12.6
0
1.6
0
6.0
0
8.0
0
8.6
0
0
0
14.4
3.8
3.2
4.8
15.6
0.8
0
0
4.7
8.3
<2.0
4.2
-------
TABLE 23 - Concluded
WATER SAMPLE ANALYSES - PERMEABLE PLUG TEST NO. 1 - MINE NO. 40-085
Date
Sampled
1-30-69
1-30-69
1-30-69
1-30-69
2- 4-69
2- 4-69
2- 4-69
2- 4-69
2- 6-69
Location
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Test Terminated
Cond.
(umhos/cm)
3010
2900
3325
3325
3000
3120
3160
3180
£H
3.6
6.6
6.8
6.7
3.7
6.9
7.0
7.0
Acidity
as CaC03
(mg/D
510
30
60
50
545
93
82
2
Alka.
as CaCOo
(mg/1)
0
204
250
252
0
254
269
267
Hardness
as CaCOs
(mg/1)
1170
1420
1680
1650
1390
1350
1690
1870
Iron
(mg/D
131.9
17.3
2.4
12.8
97.6
0.7
1.5
6.7
Sulfate
(mg/1)
1800
1280
1750
1650
1920
2000
1840
1840
Al umi num
(mg/D
14.5
< 2.0
< 2.0
2.0
13.4
0
< 2.0
< 2.0
-------
TABLE 24
FLUID FLOW DATA - PERMEABLE PLUG TEST NO. 2
ro
i
Date
5- 6-69
5- 7-69
5- 8-69
5- 9-69
5-12-69
5-14-69
5-15-69
5-23-69
5-30-69
6- 2-69
6- 4-69
6- 7-69
6-10-69
6-16-69
6-23-69
6-30-69
Total
Input
(gpm)
57.1
55.5
58.0
64.5
65.1
57.1
62.5
60.0
60.9
64.3
63.4
64.7
61.6
57.1
49.2
42.5
> Compartment
1 Compartment
1 Compartment
Over-
flow
(gpm)
36.8
32.5
39.0
41
.4
43.3
36.1
41.5
39
39
43
42
43
40
36
28
21
.0
.9
.3
.4
.7
.6
.1
.2
.5
covered
open -
covered
COMPARTMENT NO. I1
Input Dlsch Flu'
(gpm) (gpm) Pt.1
7.4 8.2
8.1
7.0
8
8
7
7
7
7
7
7
7
7
7
7
7
.4
.5
.0
.0
.0
.0
.0
.0
.0
.0
.0
.0
.0
at 36
filled
at 36
8.1
8.1
8.2
8.5
7.0
7.0
2.7
2.6
1.2
0.7
0.5
0.2
0.1
0.7
0
Inches
to top
Inches
-
12.5
16.0
29.0
31.5
-
34.3
40.8
42.3
43.8
43.8
43.8
30.5
30.0
18.0
. filled
with 17
- filled
Id Level . Inches
Pt.2 Pt.3 Pt.4 Pt.5
-
-
7.3
10.3
13.5
17.5
-
33.5
19.5
13.0
10.0
7.0
5.5
3.0
6.3
0
-
-
6.8
6.8
8.5
10.8
-
26.5
14.5
12.8
8.3
4.8
3.0
0
4.0
0
-
-
4.0
4.8
6.3
7.3
-
11.0
16.0
17.5
7.8
8.5
2.8
0
0
0
-
-
3.3
4.0
3.8
4.5
-
6.0
6.0
5.0
2.5
2.0
1.8
0
0
0
with 112 limestone.
limestone.
with 17 limestone.
COMPARTMENT NO. 2*
Input
(gpm)
6M
. *»
77
. /
6.1
7.3
6.2
7.0
7 n
i ,\j
7.0
7.0
7.0
7.0
7.0
7.0
7.0
7.0
7.0
Dlsch
(gpm)
6?
»&
7 n
/ .u
7.3
6.9
6.1
7.0
7ft
.U
7.0
6.9
4.7
3.4
7.0
7.0
3.2
0.7
4.4
Fluid Level. Inches
Pt.l Pt.2 Pt.3 Pt.4 Pt.5
1.0 3.3 4.5 3.3 3.0
3.5 5.0 6.0 5.0 5.0
7.3 8.3 8.0 7.3 4.3
11.5 12.3 11.5 10.0 6.0
17.3 20.0 16.8 18.0 11.5
35.5 36.3 35.5 29.8 16.3
40.0 40.8 39.5 32.0 18.3
41.0 42.5 41.8 33.3 19.0
40.0 42.3 42.5 37.0 20.8
43.5 43.5 43.5 37.3 21.8
42.5 40.5 43.3 37.5 32.0
20.5 21.3 20.0 20.8 11.3
26.8 24.3 22.5 42.8 25.5
COMPARTMENT NO. 3'
Input
6.5
7.2
5.9
7.5
7.1
7.0
7.0
7.0
7.0
7.0
7.0
7.0
7.0
7.0
7.0
Dlsch
6.4
7.0
7.0
7.0
6.7
7.0
7.0
7.0
7.0
7.0
7.0
7.0
5.4
1.1
0.6
Fluid Level, Inches
Pt.l Pt.2 Pt.3 Pt.4 Pt.5
.
4.0 3.8 2.8 2.3 2.3
4.5 4.5 4.8 3.3 3.0
7.8 8.0 8.5 6.5 3.8
13.3 13.5 13.0 10.0 5.8
30.0 30.3 24.5 19.3 11.3
30.5 30.5 30.3 27.5 17.5
32.0 31.3 30.5 30.0 19.0
33.5 31.3 33.3 30.5 21.0
36.8 31.8 31.0 30.5 23.3
40.0 34.0 32.0 31.0 24.5
43.0 38.0 38.3 33.8 26.3
43.5 39.0 34.0 38.8 14.5
32.3 32.3 20.0 15.3 8.0
-------
en
to
i
TAllE 8
• L U 1 B « I 0 U OUT* - MODirico r C ft M C « 8 I C PLUS TEST HO. 2
CBIPMTHtHT HO. 1
HO. 2
. 3
MM input Input
F1o» HMd Input Olictl. Fluid Ltv«1 - litellM Input Olsch. Fluid le»e» - Inches Held Input Oltch. Fig Id Level - Incho
Pete t«p») Hn.l (TO») (ap») JJJ. JJ-i Fj.3 Jjj4 Jj^J (aim) (ocm) J£j. JtJL UilISsl Ilii <'"-l (w») («") jtA Pt.Z Pt.3 ft.4 Pt.S
7- 2-4* • 24
7- W» - 24
7- 74* 17.S 24
7-21-4* 34.7 24
7-2M9 2S.7 24
7-2*-4* - X
4-18-49 41.2 M
*-2S-H 42.( II
i-28-4* M
»- 2-« M.I N
»-10-41 M
*-22-M 2t.« 24 0.2 0.2 41.5 S.O 0 0 0
t-X-4t 2».7 M 0.5 0.5 51.0 11.0 2 S.S 0
W-1V44 • T*tt
4.2 40.5 11.5 15.8 10.1 4.5
1.0 17.1 24.5 12.0 (.0 1.0
0.7 11.1 7.5 4.1 1.1 1.0 7.0 1.1 40.0 10.3 D.5 21.3 27.0
0.2 11.5 0000 7.0 0 4.5 4.5 2.0 0 0
0.2 11.0 4.0 0 0 0 7.0 1.6 11.014.3 1.0 t.O 11.0
0.4 M.5 4.0 0 0 0 7.0 0.4 44.5 11.0 11.0 12.1 39.5
0.2 15.5 0 0 0 0
TEST
OI5CONTIIDCO
7.0 0.4 27.5 13.1 12.0 t.5 4.1
7.0 0.2 18.0 (.0 8.0 4.5 2.1
7.0 0.2 18.0 11.5 9.5 t.O 1.0
7.0 0.2 16.0 11.5 11.5 8.5 6.8
7.0 0.1 11.5 11.0 11.0 9.5 7.5
24.0
27.0
».s
M.O
X.O 1.* 1.* - M.I 19.8 26.8 12.0
W.O 0.7 0.7 • 22.0 22.5 17.0 t.O
' Test Olicontlnued-
12.0
7.1 0 41.5 46.0 37.0 21.0
Enct*4 V • 2' i S' Man iu>« txumloii on COM. n. 1.
Nrforitloni pluaatd In JfltHbutton ptpt - COM. M. Z.
When op«n«dt iMttr r«n over top of rock.
F*rfor«tiont plugged In dlitrlbutton pip* - Cov. M. 2.
Knen op»«d. Mtir nn ovtr top of roe*.
Coup. «o. 2 - Mttr rions over rock to 't. S.
Inerciied Hud on Cong. No. 1.
F1o» Into Coop. Ho. t Internment due to mil raenotr
upiclty.
Teit discontinued for Cow. no. 2.
Injulled lid md pipe riser on Conp. Do. 1.
Installed punplno systea to supply siap eitensfoni.
8otn cocpertnenti no» under consUnt held »nd fle»1n».
-------
01
•fk
TABLE 26
WATER SAMPLE ANALYSES - PERMEABLE PLUG TEST NO. 2 - MINE NO. 40-085
Date
Sampled
5- 6-69
5- 6-69
5- 6-69
5- 6-69
5- 7-69
5- 7-69
5- 7-69
5- 7-69
5- 8-69
5- 8-69
5- 8-69
5- 8-69
5- 9-69
5- 9-69
5- 9-69
5- 9-69
5-10-69
5-10-69
5-10-69
5-10-69
5-11-69
5-11-69
5-11-69
5-11-69
5-12-69
5-12-69
5-12-69
5-12-69
Location
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Cond.
(umhos/cm)
2880
2700
2840
2834
2860
2720
2820
2800
2740
2640
2720
2760
2780
2680
2768
2808
2800
2740
2856
2820
2730
2674
2780
2760
2660
2700
2720
2680
£H
3.1
5.8
3.1
3.2
3.2
5.8
3.3
3.4
3.3
5.9
3.3
3.3
3.1
5.8
3.2
3.2
3.5
5.9
3.3
3.4
3.4
6.0
3.3
3.3
3.8
5.9
3.6
3.7
Acidity
as CaCOo
(mg/1)
450
213
365
335
425
200
360
300
385
335
360
350
420
155
305
310
450
135
340
305
385
140
320
280
385
175
320
275
Alka.
as CaCOs
(mg/1)
0
32
0
0
0
28
0
0
0
38
0
0
0
35
0
0
0
37
0
0
0
49
0
0
0
55
0
0
Hardness
as CaCOs
(mg/1)
1015
1400
1050
1150
1075
1305
1125
1085
965
1345
975
1030
500
1275
1115
1210
1095
1565
1190
1195
1095
1375
1190
1250
950
1450
1125
1085
Iron
(mg/1)
60.8
2.7
68.7
83.2
76.8
67.2
74.4
74.8
82.8
65.2
67.2
64.0
89.4
57.4
49.7
49.4
83.8
60.4
49.9
49.3
84.9
57.1
44.4
48.1
75.0
56.5
50.7
48.9
Sulfate
(mg/1)
1680
1855
1960
1785
1855
1820
1890
1855
1820
1925
1855
1785
1785
1750
2030
1750
1820
1820
1750
1855
1855
1785
1785
1785
1750
1785
1659
1575
Aluminum
(mg/1)
10.6
1.4
6.1
7.7
10.6
2.0
2.1
8.1
10.8
2.3
8.9
9.1
11.8
2.1
9.9
9.8
11.9
2.2
9.6
11.5
13.3
2.9
10.9
10.2
13.7
3.0
10.5
8.5
-------
TABLE 26 - Continued
WATER SAMPLE ANALYSES - PERMEABLE PLUG TEST NO. 2 - MINE NO. 40-085
CJl
en
i
Date
Sampled
5-13-69
5-13-69
5-13-69
5-13-69
5-14-69
5-14-69
5-14-69
5-14-69
5-15-69
5-15-69
5-15-69
5-15-69
5-16-69
5-16-69
5-16-69
5-16-69
5-19-69
5-19-69
5-19-69
5-19-69
5-21-69
5-21-69
5-21-69
5-21-69
5-23-69
5-23-69
5-23-69
5-23-69
Location
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Cond.
(umhos/cm)
2708
2668
2798
2718
2714
2772
2720
2650
2920
2680
2860
2840
2880
2700
2920
2848
2730
2660
2760
2700
2880
2646
2806
2640
2714
2640
2720
2740
£H
3.6
6.3
3.6
4.0
3.6
6.7
4.0
4.6
3.1
5.9
3.4
3.2
3.1
5.8
3.2
3.3
3.0
5.6
3.1
3.1
3.0
6.0
3.6
3.5
3.1
5.7
2.9
2.9
Acidity
as CaC03
(mg/ir
421
140
285
200
400
135
350
251
450
222
320
320
400
170
385
300
441
112
353
205
420
106
325
250
408
100
270
305
Alka.
as CaCOQ
(mq/1)
0
52
0
0
0
73
0
0
0
32
0
0
0
36
0
0
0
54
0
0
0
46
0
0
0
70
0
0
Hardness
as CaCOs
(mg/l)
829
2075
2715
1275
1140
1380
1150
1190
1030
1180
1095
1130
1050
1185
1095
1145
1055
1160
1125
1360
1030
1415
1225
1155
220
1520
360
1205
Iron
(mg/1)
86.3
60.6
44.7
49.6
75.2
49.3
55.6
61.8
84.7
60.8
50.6
47.4
71.4
60.7
48.9
40.9
60.7
39.3
45.4
45.9
46.3
5.4
5.7
43.8
81.4
59.2
58.5
54.3
Sulfate
(mg/1)
1785
1785
1645
1750
1715
1820
1715
1645
1750
1890
1820
1750
1050
1715
1785
1960
1820
1820
1855
1799
1680
1750
1470
1750
1680
1890
2070
1890
Aluminum
(mg/1)
13.5
3.7
10.5
8.2
13.7
3.7
8.4
8.4
14.5
0.8
10.0
7.4
12.0
0.8
10.0
7.0
12.5
2.0
9.0
8.0
12.0
7.0
7.0
6.0
12.3
2.5
6.0
7.9
-------
cji
en
TABLE 26 - Continued
WATER SAMPLE ANALYSES - PERMEABLE PLUG TEST NO. 2 - MINE NO. 40-085
Date
Sampled
5-26-69
5-26-69
5-26-69
5-26-69
5-28-69
5-28-69
5-28-69
5-28-69
5-30-69
5-30-69
5-30-69
5-30-69
6- 2-69
6- 2-69
6- 2-69
6- 2-69
6- 4-69
6- 4-69
6- 4-69
6- 4-69
6-10-69
6-10-69
6-10-69
6-10-69
6-16-69
6-16-69
6-16-69
6-16-69
Location
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Cond.
(umhos/cm)
2628
2720
2620
2768
2750
2784
2760
2880
2850
2772
2864
2788
2774
2830
2770
2876
2640
2980
2580
2760
2692
2928
2762
2824
2644
2860
2626
2666
pH
2.8
5.8
3.2
2.9
3.5
5.9
3.8
3.2
3.1
6.0
3.3
3.2
3.4
6.2
3.5
3.1
3.7
6.2
5.8
3.7
3.8
6.1
3.4
3.1
3.9
6.4
5.7
3.6
Acidity
as CaCOa
(mg/1)
425
15
220
275
450
115
270
365
426
25
230
292
425
22
210
323
398
51
162
329
525
134
430
405
465
115
171
430
Alka.
as CaCOa
(mg/1)
0
126
0
0
0
69
0
0
0
70
0
0
0
118
0
0
0
193
40
0
0
231
0
0
0
224
22
0
Hardness
as CaCOa
(mg/1)
1070
1510
1495
1455
1020
1375
1275
1145
1090
1490
1235
1220
1220
1345
1205
1170
1070
1600
1395
1170
1065
1445
1135
1140
1010
1315
1320
950
Iron
(mg/1)
81.4
55.8
69.5
58.2
25.2
51.1
55.3
54.1
85.5
51.4
52.6
61.9
95.7
44.7
62.0
66.5
74.2
117.3
59.5
67.0
99.7
24.3
92.1
81.1
95.4
4.5
66.3
80.3
Sulfate
(mg/1)
1800
1830
1830
1860
1800
1800
1830
1680
1680
1740
1680
1770
1770
2040
1980
1800
1855
1750
1785
1750
1785
1855
1855
1820
1820
1890
1855
1750
Aluminum
(mg/1)
14.4
1.3
8.7
10.1
14.6
2.7
8.1
11.7
14.3
2.0
7.9
12.7
13.1
1.3
7.6
11.4
12.4
1.3
6.4
10.7
14.6
0.8
10.3
11.6
12.7
1.4
7.9
11.1
-------
TABLE 26 - Continued
WATER SAMPLE ANALYSES - PERMEABLE PLUG TEST NO. 2 - MINE NO. 40-085
I
Ol
Date
Sampled
6-23-69
6-23-69
6-23-69
6-23-69
6-30-69
6-30-69
6-30-69
6-30-69
7- 7-69
7- 7-69
7- 7-69
7- 7-69
7-21-69
7-21-69
7-21-69
7-21-69
7-28-69
7-28-69
7-28-69
7-28-69
8- 4-69
8- 4-69
8- 4-69
8- 4-69
8-11-69
8-11-69
8-11-69
8-11-69
Location
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp . #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Cond.
(umhos/cm)
2720
2760
2860
2760
2770
2590
2710
2634
2572
2660
2540
2654
2558
2684
-
2704
2428
2476
2492
2480
2216
2342
2108
2424
2282
2408
2256
2412
JDH
3.7
5.8
6.1
6.0
3.3
6.0
3.5
6.4
3.5
5.9
3.5
6.0
3,9
6.6
-
6.6
2.9
6.6
3.2
6.9
3.4
6.4
3.5
6.7
3.3
6.4
5.7
6.8
Acidity
as CaCO?
(mg/1)
500
23
90
86
503
103
425
275
525
53
375
20
425
16
-
19
327
94
280
91
530
1070
310
1215
495
117
370
192
Alka.
as CaCOs
(mg/1)
0
132
139
174
0
61
0
96
0
118
0
163
0
193
-
175
0
190
0
219
0
162
0
209
0
199
22
246
Hardness
as CaC(h
(mg/1)
1015
1610
1500
1520
1105
138
1020
1415
930
1465
1045
1550
970
1185
-
1175
1250
1370
1160
1135
890
1205
1250
1370
159
1020
700
750
Iron
(mg/1)
105.9
43.4
49.8
73.2
98.7
84.4
73.8
73.0
103.4
70.4
74.9
56.3
88.9
23.6
-
34.8
95.7
31.7
55.6
18.4
125.4
41.9
76.5
30.8
88.6
39.4
58.6
12.1
Sulfate
(mg/1)
1820
1785
1855
1855
1505
1440
1610
1450
1410
1540
1530
1295
780
1380
-
1330
1400
1290
1260
1320
1505
1400
1435
1365
1625
1440
1615
1410
Aluminum
(mg/D
13.8
1.5
1.3
4.8
13.0
0.7
9.8
0
12.2
0
7.3
0
13.7
0
-
0
13.0
0.3
10.1
0.5
13.5
0
8.0
0.9
11.6
0.7
7.0
0.6
-------
TABLE 26 - Concluded
WATER SAMPLE ANALYSES - PERMEABLE PLUG TEST NO. 2 - MINE NO. 40-085
1
01
oo
Date
Sampled
8-18-69
8-18-69
8-18-69
8-18-69
8-25-69
8-25-69
8-25-69
8-25-69
9- 2-69
9- 2-69
9- 2-69
9-18-69
9-18-69
9-18-69
9-22-69
9-22-69
9-22-69
9-30-69
9-30-69
9-30-69
Locati on
Mine Flow
Comp. #1
Comp, #2
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Comp. #3
Mine Flow
Comp. #1
Comp. #3
Mine Flqw
. Comp. #1
Comp. #3
Mine Flow
Comp. #1
Comp. #3
Mine Flow
Comp. #1
Comp. #2
Cond.
Acidity
as CaCO-3
(umhos/cm) pH
2320
2422
2354
2462
2402
2446
1042
2446
2232
2348
1876
2660
2740
2602
2604
2722
2504
3032
2822
2702
3.2
6.5
6.4
6.9
6.6
6.5
5.0
4.8
3.2
6.5
3.3
3.2
6.7
3.6
2.9
6.8
6.0
3.0
6.7
6.3
(mg/1)
395
0
0
37
445
43
0
176
393
98
341
417
26
277
505
3
91
293
21
5
Alka.
as CaCOs
(mg/1)
0
184
74
227
0
189
126
204
0
192
0
0
233
0
0
250
51
0
161
94
Hardness
as CaCOs
(mg/1)
990
1270
1270
1325
915
1340
87
1230
1040
1380
1080
1160
1635
1295
850
1355
1350
1160
1550
1250
Iron
(mg/1)
96.8
43.5
25.9
32.3
98.2
41.2
0.5
32.5
93.3
47.0
81.6
89.6
17.8
70.0
91.2
18.0
70.0
106.0
-
87.3
Sulfate
(mg/1)
1260
1530
1560
1470
1500
1470
40
1500
1500
1470
1500
1410
1295
1500
1440
1645
1530
1520
—
1680
Aluminum
(mg/1)
7.9
0.7
4.8
1.2
11.5
1.5
1.7
2.1
13.9
0
13.7
12.6
2.5
6.4
12.2
0.2
1.0
8.2
—
0.6
-------
en
10
TABLE 27
FLUID DATA AND WATER ANALYSES
PERMEABLE PLUG TEST NO. 3 - MINE NO. 62-008
Date
Sampl ed
9-21-68
6- 2-69
6-10-69
6-16-69
6-23-69
6-30-69
7- 7-69
7-21-69
7-28-69
8- 4-69
8-11-69
8-18-69
8-25-69
9- 2-69
9-16-69
9-22-69
9-29-69
10- 6-69
10-13-69
Flow
(gpm)
3.0
3.3
1.8
2.5
2.4
2.3
2.4
3.4
2.4
3.4
3.0
2.4
4.0
3.2
2.2
2.2
3.0
2.4
Head
(in.)
-
_
32. 01
33.0
36.3
39.5
42.3
42.0
42.5
44.0
44.5
46.8
47.3
49.3
49.5
49.8
50.0
49.8
Cond.
(umhos/cm)
4700
-
4222
4186
4200
4160
4240
4202
3950
3886
3700
3080
3820
3724
4322
4242
4220
-
-
EH_
4.3
-
3.0
6.3
6.5
6.9
6.4
6.9
6.7
6.8
6.6
6.7
6.7
6.5
6.6
6.5
6.5
-
-
Acidity
as CaCOo
(«ng/i)
100
-
300
7
61
58
200
n
108
315
161
39
286
222
84
4
12
-
-
Alka.
as CaCOo
(mg/ir
0
-
0
255
265
301
284
248
234
260
253
232
234
219
182
200
181
-
-
Hardness
as CaCOo
(mg/ir
1490
-
1445
1595
1630
1005
1645
1150
1830
1730
1110
1490
1505
1840
1250
1205
1935
-
-
Iron
(mg/1)
36.5
-
75.3
13.2
7.2
4.2
6.4
28.7
31.6
26.8
24.0
45.1
33.5
55.1
57.2
66.4
72.3
-
-
Sulfate
(mg/D
1690
-
2750
2500
2850
2550
2200
2200
2200
2350
2700
2760
2610
2340
2200
2650
2500
-
—
Aluminum
(tng/1 )
3.2
-
5.0
0
0
0
0
0
0.6
0.1
0.4
1.2
1.7
0
2.9
0.2
-
—
—
Built plug of graded limestone in Opening No. 3 - Mine No. 62-008.
-------
TABLE 28
SCREEN ANALYSES
LIMESTONE USED IN MINE NO. RT5-21
AASHO #67
U.S. Standard
Sieve Mesh Size
PassedRetained
3/4
" *
3/8
"*
3/4
"*
3/8
"*
Per Cent
Retained
78
97
AASHO #83
U.S. Standard
Sieve Mesh Size
PassedRetained
1/2"* 3/8"*
3/8
"*
8
Per Cent
Retai ned
11
76
97
8
99
8
16
99
Elkins Limestone Co., Inc. - Elkins, West Virginia
Equivalent to West Virginia No. 7 Limestone.
Equivalent to West Virginia No.12 Limestone.
Size of opening in inches.
-160-
-------
TABLE 29
MONITORING DATA
SEAL NO. 1 - MINE NO. RT5-2
Flow Rate
Date (ft) (gpm)
7-22-69 - 58.5 Bulkhead seal completed 7-20-69 -
drain valve still open.
7-22-69 - 0 Drain valv.e closed at 2:40 p.m.
7-28-69 2.92 - Small leakage-right-hand side of
mine.
7-29-69 3.22 - Leakage on right-hand side in-
creasing - remedial work necessary.
9-15-69 - 20.1 Remedial work completed by sealing
adjacent opening. All drain lines
open.
Drain valves closed.
9-17-69
9-18-69
9-19-69
9-22-69
9-25-69
9-26-69
9-29-69
9-30-69
10- 1-69
10- 2-69
10- 3-69
10- 6-69
10- 7-69
10- 8-69
10- 9-69
10-13-69
10-14-69
10-15-69
0.81
1.20
1.68
2.62
3.25
3.36
3.60
3.62
3.64
3.66
3.70
3.70
3.72
3.72
3.74
3.78
3.78
3.78
-161-
-------
TABLE 30
MONITORING DATA
SEAL NO. 2 - MINE NO. RT5-2
Date
9-17-69
9-18-69
9-19-69
9-22-69
9-25-69
9-26-69
9-29-69
9-30-69
10- 1-69
10- 2-69
10- 3-69
10- 6-69
10- 7-69
10- 8-69
10- 9-69
10-13-69
10-14-69
10-15-69
Head
(ft)
2.58
2.92
3.25
3.75
4.90
5.00
5.16
5.26
5.28
5.32
5.34
5.36
5.38
5.39
5.40
5.44
5.44
5.44
Flow Rate
(gpm)
1.2
3.4
i
25.0
21.0
16.2
12.1
9.0
9.0
16.2
16.2
16.2
9.0
4.9
3.6
3.6
3.6
3.6
1 No flow taken due to construction.
-162-
-------
I
CO
TABLE 31
WATER ANALYSES - MINE NO. RT5-2
Alka. Hardness
Date
Sampled
9-22-69
10- 1-69
10- 1-69
10- 1-69
10-13-69
10-13-69
10-13-69
Cond. as CaCOs
Location (umhos/cm) pH (mg/1)
Opening
Opening
Opening
Opening
Opening
Opening
Opening
2-B1
I2
2-A3
2-B
1
2-A
2-B
1484
1400
1400
1400
1400
1220
1160
6.2
2.9
2.9
6.3
3.1
3.2
5.6
103
696
557
65
765
512
0
as CaCOs
(mg/1)
77
0
0
0
0
0
174
as CaCOs
(mg/1)
895
513
544
1279
657
663
1450
Iron
(mg/1)
36.
104.
114.
57.
282.
180.
36.
9
0
0
0
0
0
0
Sulfate
(mg/1)
780
1152
988
940
916
868
783
Aluminum
(mg/1)
5
35
38
0
53
40
3
i
Discharge from Opening No. 2 through limestone seal.
2 Water from behind silicate cement bulkhead in Opening No. 1.
3 Water from behind limestone seal in Opening No. 2.
-------
BIBLIOGRAPHIC: Halliburton Company, "New Mine Seal-
ing Techniques for Water Pollution Abatement"
FWPCA Publication No. 14010 DM0
ABSTRACT: The purpose of this project was to de-
velop and field test new concepts applicable to
abatement of acid mine water pollution. Labo-
ratory research determined proper materials,
equipment and techniques for constructing mine
seals. Two new processes were developed. One
involved a technique of placing a plug of
graded limestone aggregate in a mine drift to
neutralize acid mine water discharge until
a seal was effected. In the process precipita-
tion of iron hydroxide gradually closed the
pores. A pneumatic conveying technique per-
mitted placement of 550 pounds per minute of
aggregate into a mine drift. The second pro-
cess consisted of remotely constructing a mine
seal including rear and front bulkheads of
self-supporting, quick-setting sodium silicate
cement specifically developed for this appli-
cation. A filler material of expansive cement
was used between the bulkheads to complete the
KEY WORDS:
Research and Develop-
ment
Acid Mine Water
Pollution Abatement
New Mine Sealing
Concepts
Self-Supporting
Buikheads
Remote Bulkhead
Construction
Permeable Limestone
Seals
Acid Neutralization
in Situ
-------
seal. Field testing in West Virginia substan- Pneumatic conveying
tiated the feasibility of both processes when Techniques
two aggregate and two bulkhead type seals were
placed in abandoned mines with drainage flows Mine Sealing Costs
up to 58 gallons per minute. Cost of the
seals are reported. This report was submitted Grouting
in partial fulfillment of Contract No.
14-12-453 between the Federal Water Pollution
Control Administration and the Halliburton
Company.
* U.S. GOVERNMENT PRINTING OFFICE : 1970 O - 404-763
------- |