EPA/600/R-05/071
April 2005
Mine Waste Technology Program
Underground Mine Source Control
Demonstration Project
by:
Lynn McCloskey
MSB Technology Applications, Inc.
Mike Mansfield Advanced Technology Center
Butte, Montana 59702
Under Contract No. DE-AC09-96EW96405
Through EPA lAGNo. DW89938870-01-1
Diana Bless, EPA Program Manager
Sustainable Technology Division
National Risk Management Research Laboratory
Cincinnati, Ohio 45268
This study was conducted in cooperation with
U.S. Department of Energy
Savannah River Operations Office
Aiken, South Carolina 29802
National Risk Management Research Laboratory
Office of Research and Development
U.S. Environmental Protection Agency
Cincinnati, Ohio 45268
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Notice
The U.S. Environmental Protection Agency through its Office of Research and Development funded the
research described here under IAG DW89938870-01-1 through the Department of Energy (DOE)
Contract DE-AC09-96EW96405. It has been subjected to the Agency's peer and administrative review
and has been cleared for publication as an EPA document. Reference herein to any specific commercial
product, process, or service by trade name, trademark, manufacturer, or otherwise, does not necessarily
constitute or imply its endorsement or recommendation. The views and opinions of authors expressed
herein do not necessarily state or reflect those of the U.S. EPA or DOE, or any agency thereof.
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Foreword
The U.S. Environmental Protection Agency is charged by Congress with protecting the Nation's land, air,
and water resources. Under a mandate of national environmental laws, the Agency strives to formulate
and implement actions leading to a compatible balance between human activities and the ability of natural
systems to support and nurture life. To meet this mandate, EPA's research program is providing data and
technical support for solving environmental problems today and building a science knowledge base
necessary to manage our ecological resources wisely, understand how pollutants affect our health, and
prevent or reduce environmental risks in the future.
The National Risk Management Research Laboratory is the Agency's center for investigation of
technological and management approaches for preventing and reducing risks from pollution that threaten
human health and the environment. The focus of the Laboratory's research program is on methods and
their cost effectiveness for prevention and control of pollution to air, land, water, and subsurface
resources; protection of water quality in public water systems; remediation of contaminated sites,
sediments, and ground water; prevention and control of indoor air pollution; and restoration of
ecosystems. The NRMRL collaborates with both public and private-sector partners to foster technologies
that reduce the cost of compliance and to anticipate emerging problems. NRMRL's research provides
solutions to environmental problems by developing and promoting technologies that protect and improve
the environment; advancing scientific and engineering information to support regulatory and policy
decisions; and providing the technical support and information transfer to ensure implementation of
environmental regulations and strategies at the national, state, and community levels.
This publication has been produced as part of the Laboratory's strategic long-term research plan. It is
published and made available by EPA's Office of Research and Development to assist the user
community and to link researchers with their clients.
Sally Gutierrez, Director
National Risk Management Research Laboratory
in
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Abstract
This report presents results of the Mine Waste Technology Program Activity III, Project 8, Underground
Mine Source Control Demonstration Project implemented and funded by the U. S. Environmental
Protection Agency (EPA) and jointly administered by EPA and the U. S. Department of Energy (DOE).
Project 8 addresses EPA's technical issue of Mobile Toxic Constituents - Water - through a field
demonstration at a remote, inactive underground mine.
This project was undertaken to demonstrate the feasibility of injecting a water-activated, expansive,
flexible, and closed-celled source control material (grout) into a rock fracture system to reduce or
eliminate the flow of acid mine drainage (AMD) into the underground workings of an abandoned mine.
The Miller Mine, located in the Big Belt Mountains of Broadwater County, Montana, was selected for the
field demonstration.
Grout injection was completed in two phases to reduce or eliminate AMD generated when groundwater
contacts sulfide mineralization associated with the Miller Mine Reverse Fault and associated fractures.
Phase I included drilling, coring, and grouting 10 approximately 624-feet drill holes in Precambrian
sedimentary and Tertiary igneous rocks using core and Jackleg drilling methods. Results of Phase I work
indicate that fracture flow was significantly reduced by approximately 77 +/- 5%, and the metals loading
(particularly for arsenic, cadmium, copper, nickel, aluminum, and lead) was reduced by at least 80%.
Phase II grout injection included using Jackleg drilling and downstage grouting methods (approximately
400 feet) completed to seal rock fractures with observed flow that intercepted underground workings.
Phase II work further reduced metals loading (zinc, iron, and copper) into the underground workings.
While the dissolved metal concentrations at the portal were not reduced, the mass loading of the major
metals (iron, zinc, and aluminum) were reduced by an order of magnitude. This is the direct result of
grout injection into the fracture system. The grouting reduced fracture flow and the metals load resulting
from the diversion of fracture flow away from mineralized areas and encapsulation of the sulfide
mineralization.
IV
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Contents
Page
Notice ii
Foreword iii
Abstract iv
Contents v
Figures viii
Tables ix
Acronyms and Abbreviations x
Acknowledgments xi
Executive Summary ES-1
1. INTRODUCTION 1
1.1 Project Description 1
1.2 Technology Background 2
1.2.1 Technology Background 2
1.2.2 Project Objectives and Scope of Work 2
1.2.3 Technology Criteria 3
1.2.4 Demonstration History of Field Emplacements 3
2. PREELECTION SITE CHARACTERIZATION SUMMARY 8
2.1 General Information, Surveys, and Observations 8
2.2 General Site Geology 8
2.3 Miller Mine Hydrogeology 8
2.3.1 Underground Mine Flow Regime 9
2.3.2 AMD Mine Discharge and Water Quality 9
2.4 Geophysical Survey 10
3. GROUT INJECTION PROCEDURES 13
3.1 Introduction 13
3.2 October 1999 Core Drilling and Recovery 13
3.2.1 Recovery of Orientated Core 13
3.3 Water Injection and Dye Tests 14
3.4 Grout Formulations and Emplacement 14
3.5 April 2001 - Wl Drift Drilling and Grouting 14
3.6 Geophysical Investigation Methods 15
4. PHASE I - FIELD EMPLACEMENT RESULTS 16
4.1 Preliminary Design (1999) 16
4.1.1 Design Segment 1 - Grout Injection 17
4.1.2 Design Segment 2 - Grout Injection 17
4.1.3 Design Segment 3 - Grout Injection 17
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Contents (cont'd)
Page
4.2 Rock-Core Drilling and Recovery 17
4.2.1 Downhole Camera Survey 18
4.2.2 Core Hole Location Survey 18
4.2.3 Geology - Recovered Rock Core 18
4.2.4 Structure 19
4.2.5 Fracture/Shear System - Mine Map 19
4.2.6 Fracture/Shear Systems - Rock Core 19
4.2.7 Observed Water-Bearing Fractures 19
4.3 Water/Dye Injection Tests 21
4.3.1 Lugeon Water Injection Testing 21
4.3.2 Lugeon Fracture Hydraulic Conductivity Results 21
4.4 Phase I - Grout Emplacement 22
4.4.1 Core Hole Fracture Cross Communication 23
5. PHASE II - FIELD EMPLACEMENT 30
5.1 Preliminary Design 30
5.1.1 Geology and Drift Segments (A, B, C, D, and L13) 30
5.2 Summary of April 2001 Grout Emplacement 31
5.2.1 Grout Emplacement 32
6. LONG-TERM MONITORING AND MINE MAINTENANCE RESULTS 37
6.1 Mine Maintenance 37
6.2 Monitoring History and Methods 37
6.3 Long-Term Monitoring Results 37
6.4 Water Quality Results at the Miller Mine 38
6.4.1 Water Quality and Metals Loading at Wl 39
6.4.2 W4 Miller Mine Portal Water Quality and Metals Loading 39
7. SUMMARY OF QUALITY ASSURANCE ACTIVITIES 46
7.1 Background 46
7.2 Project Reviews 46
7.3 Data Evaluation 46
7.3.1 Analytical Evaluation 46
7.3.2 Program Evaluation 46
7.4 Summary 47
VI
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Contents (cont'd)
Page
8. CONCLUSIONS AND RECOMMENDATIONS 50
8.1 Lessons Learned 51
9. REFERENCES 52
Appendix A: Miller Mine Maps A-
AppendixB: Miller Mine Water Quality Data B-
AppendixC: 1999 Core Logs C-
AppendixD: Geophysical Reports fromNETL D-
Appendix E: Water Injection Test Results from October 1999 (Lugeon Values for Boreholes) E-
Appendix F: Flow Calculations and Statistical Analysis F-
Note: Appendices A through F are available upon request from the MSE MWTP Program Manager.
Please refer to document number MWTP-258. Email: mwtp@mse-ta.com. Phone: (406)494-7100.
vn
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Figures
Page
1-1. Miller Mine Map 5
1-2. Generalized Site Map of the Miller Mine 6
1-3. Photo of the Miller Mine Upper and Lower Workings 6
1-4. Photo of the Miller Mine Lower Adit Discharge Prior to Implementation of the Source Control
Technology 7
1-5. Areas of AMD at the Miller Mine Prior to the Application of the Technology 7
2-1. Initial Miller Mine Underground Workings Plan Map Showing the Weir, Flume, and Sample
Port Locations W1,W2,W3, and W4 11
3-1. The Drilling/Grouting Subcontractor Placing an Inflatable Packer System into the Grout Hole
for Dye and Water Injection Testing 15
3-2. Photograph of the Grout Pump, Tanks, and Flowmeter System for the Stage I Grout Injection
in 1999 15
4-1. Plot of the Average Lugeon and Fracture Density for Grout Hole Ml 25
4-2. Plot of the Average Lugeon and Fracture Density for Grout Hole M2 25
4-3. Plot of the Average Lugeon and Fracture Density for Grout Hole M3 26
5-1. Stabilizing the Underground Mine Working of the Miller Mine with Roof Bolts and Screen
using a Gardner-Denver Model 83 Jackleg Drill 33
5-2. Kabota Tractor that Transported the Grout and Grout Equipment into the Abandoned
Underground Workings of the Miller Mine 33
5-3. Iron Oxide Stalactites in the Wl Drift, which were Areas where Flow Through the Fractures
Required Grouting 34
6-1. Inflow (Wl, W2, W3 Added Together) and Outflow (W4) Measurements Compared over Time
to Evaluate Losses and Fluctuations due to Grout Injection 40
6-2. Miller Mine Flow Measurement Taken at the Respective Locations Throughout the Mining System.
Note that the Miller Mine Became Unstable, and Measurements for Wl, W2, and W3 were
Discontinued in November 2001 40
6-3. Miller Mine Water Quality Results for the Wl Drift, the Area where Most of the Dissolved Metals
were Originating. The Water Quality Results Reflect the Dissolved Metals Concentrations at
Wl Micrograms per Liter 41
6-4. Average Percent Reduction for Metals Loading at the Wl Drift Prior and after each Phase of
Grouting 41
6-5. Metals Loading at the Wl Drift Sample Port Located in the Underground Workings of the
Miller Mine 42
6-6. Water Quality at the Miller Mine Portal (W4 Sample Port) 42
6-7. The Average Percent Reduction in Metals Loading at the Miller Mine Portal after Grout
Injection 43
6-8. Metals Loading for the Miller Mine Portal (Sampling Results from Port, W4) 43
Vlll
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Tables
Page
2-1. Background Water Quality for the Miller Mine Prior to Field Emplacement of the Source
Control Technology 12
4-1. Rock Core and Drill Data 26
4-2. Comparison of Downhole Camera Survey Data 27
4-3. Survey - Drill Point Location 27
4-4. Water-Bearing Fractures (Intervals)/Communication 27
4-5. Average Fracture Hydraulic Conductivity Values (cm/sec) from Water Tests (K); Based on
Lugeon Values 28
4-6. Grout Properties 28
4-7. Grout Injection Data 28
4-8. Water Test/Grout Communication Between Core Holes 29
5-1. Summary of the Grout Injection Sequence at the Miller Mine, W-l Drift 35
6-1. Wl Drift Recorded Flow Data Taken for the Duration of the Technology Demonstration 44
6-2. The Volume of Flow from the Miller Mine, Drift Wl, Prior to and after each Phase of
Grouting on a Gallons per Day Basis 44
6-3. Flow Reductions at the Wl Drift 45
6-4. Wl Drift Flow Reductions with Error Ranges 45
6-5. Water Quality Results for the Miller Mine after Both Phases of Field Emplacement of the
Source Control Technology 45
7-1. QA Objectives for Accuracy, Precision, MDL, and Completeness 48
7-2. Instrument Detection Limits (IDLs) for ICP Analysis of Dissolved Metals 48
7-3. Summary of Flagged Data for Activity III, Project 8 49
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Acronyms and Abbreviations
Ag silver
Al aluminum
AMD acid mine drainage
As arsenic
Cd cadmium
cm/s centimeter per second
cps centipoises
Cu copper
DOE U.S. Department of Energy
EPA U.S. Environmental Protection Agency
Fe iron
ft foot
gpm gallons per minute
HA Hydro Active
ID inside diameter
ICP-ES inductively coupled plasma emission spectrometer
IDL instrument detection limit
Ib/d pound per day
MCL maximum contaminant level
MDL method detection limit
Mg magnesium
mg/L milligrams per liter
mL/min milliliter/minute
MMRF Miller Mine Reverse Fault
Mn manganese
MSE MSE Technology Applications, Inc.
MWTP Mine Waste Technology Program
Ni nickel
OD outside diameter
Pb lead
ppm parts per million
psi pounds per square inch
QA quality assurance
QAPP quality assurance project plan
QC quality control
RPD relative percent difference
RQD rock quality designation
TD total depth
Tqd tertiary quartz diorite
VLF very-low frequency
Zn zinc
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Acknowledgments
This document was prepared by MSB Technology Applications, Inc. (MSB) for the U.S. Environmental
Protection Agency's (EPA) Mine Waste Technology Program (MWTP) and the U.S. Department of
Energy's (DOE) Savannah River Operations Office. Ms. Diana Bless is EPA's MWTP Program
Manager, while Mr. Gene Ashby is DOE's Technical Program Officer. Ms. Helen Joyce is MSB's
MWTP Program Manager.
XI
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Executive Summary
The Mine Waste Technology Program (MWTP), Activity III, Project 8, Underground Mine Source
Control Demonstration Project was implemented and funded by the U.S. Environmental Protection
Agency (EPA) and jointly administered by EPA and the U.S. Department of Energy (DOE). Project 8
addresses EPA's technical issue of Mobile Toxic Constituents - Water - through a field demonstration at
a remote, inactive, underground mine. The Underground Mine Source Control Demonstration Project
was performed to demonstrate the feasibility of injecting an innovative, nonrigid, source control material
into the fractured rock system at an abandoned underground mine to reduce and/or eliminate the influx of
acid mine drainage (AMD) into the underground mine workings.
In 1998, the MWTP selected the Miller Mine as a demonstration site for the field implementation and
evaluation of an underground mine source control technology. Adit discharge ranged between 5 and
14.5 gallons per minute (gpm) and contained lead levels that exceeded the National Primary Drinking
Water Regulations; aluminum, iron, manganese, pH, and sulfate at levels that exceeded the National
Secondary Drinking Water Regulations; and nickel that exceeded the aquatic life standard for fresh water.
The Miller Mine discharge is used as a source of drinking water for cattle and wildlife in the area
immediately surrounding the Miller Mine.
Site characterization, source control materials testing/evaluation, and two phases of field source control
emplacement were performed to define the optimal material and material application method for the
demonstration. An expandable, closed-cell polyurethane grout material was selected. This material was
designed and developed to act as a water-stop or water barrier system and to seal subsurface structures
such as dams, tunnels, and sewers. This demonstration used Hydro Active Combi grout manufactured by
De Neef, Inc. to seal the fractured rock system that was acting as a conduit for the discharge of AMD into
the Miller Mine subsurface workings.
Site hydrogeological and geochemical conditions were monitored during the entire duration of the
project. The main lower level of the Miller Mine consists of a 600-foot drainage/haul tunnel that forks
approximately 70 feet from the end of the workings. During site characterization, it was determined that
the main sources of influx occurred in the two drifts that fork from the main tunnel. The flow in one drift
(W2), which contains a winze (a vertical workings starting in the floor of the workings), was fairly clean,
containing minimal metals contamination and flows of approximately 1.5-gpm discharge from the drift.
Before grouting, the second drift (Wl drift) contributed flows up to a maximum of 14.5 gpm, which
exceeded most of the water quality regulatory levels. The main geologic structure controlling the flow at
the end of the main workings was the Miller Mine Reverse Fault and the associated fracture system.
Phase I field source control technology injection was performed by core drilling grout injection holes
over the top of the Wl drift and perpendicular to the fracture system to reduce the inflow of water. As a
result of the Phase I grout injection, flow was reduced from the Wl drift by an average of approximately
77 +/- 5% over 10 months.
Equipment used during the Phase I grout injection was too large to use inside the Wl drift. After the
initial grout injection, water still continued to seep from the walls through fractures in the Wl drift,
especially on the northeastern wall of the underground mine workings. A decision was made to perform
Phase II work by grouting small seeps flowing from close fractures in the mine wall to determine if the
flow from the Wl drift could be further reduced or eliminated. The seeps were not eliminated, and the
ES-1
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influx through the small fractures migrated to different fractures after grouting was performed. The
maximum reduction in flow after both injections was approximately 77 +/- 5%. From long-term
monitoring results taken after grouting, the maximum recorded total flow from the Miller Mine adit was
approximately 3 gpm, where 47% of the flow was from the W2 drift and 53% of the flow was from the
Wl drift. Due to the reduction in flow from the Wl drift, the average percent reduction of dissolved
metals loading over the duration of the project and the metals loading rates were significantly reduced.
Average dissolved metal loading reductions of greater than 80% were obtained for cadmium, aluminum,
zinc, and iron. Reductions of greater than 50% were obtained for manganese, lead, nickel, and copper.
Iron loading, for example, was reduced from 7.5 pounds (lb)/day to 0.12 Ib/day at sample port Wl.
As a result of the technology application, only iron, lead, and manganese remain slightly above the
National Secondary Drinking Water Regulation levels at 0.4, 0.04, and 3.26 parts per billion,
respectively. As with the flows, these concentrations are much less than the original concentrations. An
approximately 50% reduction in metals loading was achieved for all of the metals analyzed; for zinc and
iron, the reduction was greater than 90%.
Overall, the application of the technology reduced the flow and the metals loading from the Miller Mine.
However, most of the reduction in flow was achieved and resulted from the initial grout injection
performed in 1999. The second grout application did not further reduce flows because the small,
interconnected fractures controlling the seeps/flows into the underground workings were difficult to
grout. Also, the flow would shift from one location to another because the rock was sheared and very
weak. The main positive result from the second grout injection was the reduction of metals loading at
sample point Wl and, to a lesser extent, at sample point W4 when comparing data from analogous
seasons.
ES-2
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1. Introduction
1.1 Project Description
Mine Waste Technology Program (MWTP)
Activity III, Project 8, Underground Mine Source
Control Demonstration Project was funded by the
U.S. Environmental Protection Agency (EPA) and
jointly administered by EPA and the U.S.
Department of Energy (DOE) through an
Interagency Agreement. EPA contracted MSE
Technology Applications, Inc. (MSE) through the
MWTP to develop and evaluate a source control
grouting technology that could be applied at
abandoned mines to seal flows into the
underground mine workings.
The objective of MWTP Activity III, Project 8
was to demonstrate the feasibility of source
control materials for hydrogeological control to
reduce water influx and minimize the production
of acid mine drainage (AMD) at a nonferrous
metal mine. The Miller Mine, located in the
Confederate Mining District in Broadwater
County, Montana, was the mine selected for
implementation of the technology. The source
control material selected for the demonstration
was Hydro Active (HA) Combi grout, a closed-
celled, expandable polyurethane grout that is
flexible and can be injected under cold and
submerged conditions. Details concerning the
material selection process are provided in the
MWTP Activity III, Project 8, Phase I, Site
Characterization and Materials Testing Report
(Ref 1).
Application of this technology involved injecting
the polyurethane grout into the rock-fracture
system associated with the underground mine
workings, thereby reducing the amount of ground
and surface water infiltrating the mine workings
and system. Reducing water inflow into the
underground mine workings was expected to
reduce the volume of impacted water discharging
from the lower mine adit. However, these
activities also have the potential to improve
surface water quality downstream from the mine.
The Miller Mine site, selected for this
demonstration project, is located approximately 20
miles north of Townsend, Montana. Presently,
slightly acidic waters containing elevated levels of
heavy metals discharge from the mine adit directly
into Greenhorn Gulch.
The Underground Mine Source Control
Demonstration Project consisted of three major
phases: (1) site characterization of the Miller
Mine, (2) materials testing, and (3) field
demonstration of the selected technology
including verification monitoring and technology
evaluation. This document is the final report and
will address Phase III, Field Emplacement.
However, all pertinent information pertaining to
field emplacement that was obtained during other
phases of the project will be addressed in this
document. Information from all phases of the
project was used to evaluate the effectiveness of
the technology.
Work associated with Phase I included
characterizing the Miller Mine site. The purpose
of Phase I was to define characteristics of the mine
site prior to technology application, including:
- sources of water infiltrating the underground
workings;
- flow rates and water quality within the
underground mine workings;
- the hydrogeological system of the mine site;
- historical, mineralogical, and structural
geology of the mine system; and
- hydraulic connections within the underground
mine.
Phase II involved performing materials testing of
approximately 50 source control materials to
determine which material would be viable for
application during Phase III. Phase II testing was
performed at MSB's testing facility and at IT
Geotechnical Laboratory, Inc.
Phase III activities that are addressed in this report
are given below.
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The site description and background
information are presented in Section 1.
A preinjection site characterization summary
is presented in Section 2.
The preliminary grout injection procedures
and requirements are provided in Section 3.
The general approach for the 1999 technology
field emplacement and investigation is
presented in Section 4.
The general approach used for the 2001
technology field emplacement and
investigation is presented in Section 5.
. Monitoring results and evaluation for the
entire duration of the project are provided in
Section 6.
A summary of quality assurance project plan
(QAPP) activities is provided in Section 7.
The conclusions derived from the field
program, previous work performed, and
recommendations for future projects of this
type are in Section 8.
A list of references used in the document is
recorded in Section 9.
1.2 Technology Background
1.2.1 Technology Background
The Miller Mine is an abandoned gold mine
located in steep terrain on the western slope of the
Big Belt Mountains in Broadwater County,
Montana (SW1/4, SE1/4, Sec. 13, T10N, R2E)
between Greenhorn Gulch and Montana Gulch
(Figure 1-1). Mine property is on the Superior
Claim in the Confederate Mining District
approximately 20 miles northeast of Townsend,
Montana, at an altitude of 6,400 feet (ft). The
mine includes two working levels, i.e., an upper
and lower mine workings (Figures 1-2, Figure 1-3,
and Appendix A, Plate 1). The lower adit
drainage is slightly acidic while the upper
workings are generally dry (Figure 1-3). The
discharge from the Miller Mine flows into
Greenhorn Gulch, a tributary of the creek in
Confederate Gulch, which eventually empties into
Canyon Ferry Reservoir (Figure 1-4).
Mine hazards at the Miller Mine site include two
open adits, several waste dumps, numerous
structures in various stages of disrepair, and
discarded equipment. Approximately
12,000 cubic yards of disturbed material and waste
rock are associated with the Miller Mine site
(Figure 1-3.)
1.2.1.1 Site History
Lode claims in the Confederate Gulch area were
first staked by Henry O. Miller in 1893. Mining
continued at the Miller Mine until the early 1940s
with the Miller Mine being the largest lode gold
producer in the area. Approximately 4,000 ounces
of gold ($80,000) were produced from the Miller
Mine between 1910 and 1930. The mine was
reopened after World War II and operated
intermittently until the early 1960s. Except for
exploratory drilling in recent years, very little
activity or mining activity has been recorded in the
area since the early 1960s.
1.2.1.2 Physiography
Terrain around and in the vicinity of the Miller
Mine is steep and slightly wooded with some
vegetation. Narrow, steep, and unpaved roads
provide vehicle access to most areas of the mine.
Winter access to the site is difficult due to deep
snow and steep terrain, which impedes sampling
efforts outside the mine during the winter months
(Figure 1-3). During the winter months, the Miller
Mine discharge freezes approximately 50 ft in
from where the discharge exits the mine portal.
1.2.2 Project Objectives and Scope of Work
Most waters in the vicinity of the Miller Mine
have a pH close to neutral and do not carry a large
percentage of metals or suspended solids before
infiltrating the subsurface fracture systems
associated with underground mine workings. The
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AMD forms when infiltrating water contacts an
oxidized sulfide ore zone as it travels through the
associated fracture system. Acid mine drainage
typically is caused by the oxidation of sulfide
mineralization. Especially problematic is when
iron disulfide (pyrite) comes in contact with water.
This chemical reaction results in increased acidity
(lower pH) of the water and increased mobility of
metals in water.
The objective of the Phase III field demonstration
emplacement was to inject an innovative source
control material into the rock fracture system and
reduce or eliminate AMD by limiting the
infiltration of water through hydraulic connections
associated with the fracture system at the Miller
Mine. The source control material was injected in
order to fill and seal fractures that acted as
hydraulic conduits, to reduce the amount of water
infiltrating the underground workings, to prevent
water from coming in contact with oxidized host
rocks, and ultimately, to reduce or eliminate the
generation of AMD and mobilization of heavy
metals. Figure 1-5 provides a visual depiction of
areas of AMD at the Miller Mine prior to the
application of the technology.
This project demonstrated the effectiveness of an
experimental, closed-celled, expandable,
polyurethane grout as the innovative source
control material. The grout was injected into the
localized rock-fracture system to seal and reduce
water flow into sections of the underground mine
workings. The adit discharge and the discharge
from each drift was continuously monitored before
and after grout injection to help determine if the
grout had sealed the localized fracture system and
reduced the AMD in the mine.
1.2.3 Technology Criteria
Criteria for the success of the polyurethane
grouting technology was established to define and
measure the degree of success the technology
application was able to achieve. Water-injection
field tests were performed prior to establishing the
hydraulic baseline conditions at the site before
grouting. Postgrouting tests were performed, and
results were compared to the baseline conditions
to determine the performance of the polyurethane
grout with regard to hydraulic sealing of the
fracture system. The primary objective used to
define the success of the project is noted below
and is also addressed in the QAPP (Ref 2). The
specific Phase III objective was to show that the
injection of the polyurethane grout reduced the
cumulative volumetric flow by 95% at sample port
Wl (a 60-degree trapezoidal flume). The initial
volumetric flow at Wl was determined using flow
data from the 10 months prior to grout
emplacement, and the final volumetric flow was
determined using the flow data from the
10 months following grout emplacement
(Appendix A, Plate 1).
Because of seasonal variations in the flow at the
Wl flume (i.e., high spring flow and low winter
flow), the achievement of the objective was also
evaluated on a seasonal basis. Cumulative flows
for each 2-month period before grouting were
compared to the corresponding 2-month period
after grouting to show the near-term effect of the
grout emplacement.
A secondary criteria for the success of Phase III
included improving the quality of water exiting the
Miller Mine adit by decreasing dissolved metals
concentrations and increasing the pH.
Although the main objective of the demonstration
was to control point-source influx, the evaluation
of the project's success also included the
feasibility, cost-effectiveness, and flexibility of the
technology to be used in other situations and other
applications.
1.2.4 Demonstration History of Field
Emplacements
Site characterization of the Miller Mine began in
August 1998 and was completed October 1999.
Fieldwork consisted of mine mapping, core and
borehole drilling, water injection testing, and grout
injection. Technology emplacement fieldwork
was performed in two stages.
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Stage I - Field Emplacement 1999 - The initial
technology emplacement began during the week of
September 13, 1999, and was completed by
October 21,1999. For design and characterization
prior to and during the emplacement, numerous
water injection and dye tests were performed on
several of the holes to determine the hydraulic
conductivity of the fracture system and to help
predict which fractures influenced the rock
fracture system the most during grout injection.
Fieldwork included drilling, coring, and grouting
10 core holes or drill holes (M1-M9 and MJ10)
(Appendix A, Plate 2). A total of 624 ft of drilling
was completed during this period that included
617 ft of coring. Injection grouting was associated
with core holes Ml through M9 located between
Wl and W2 in the lower working level of the
mine. Borehole JK10 (7 ft in depth) was not core-
drilled and is located northeast of Wl (Appendix
A, Plate 2).
Stage II - Field Emplacement 2001 - The second
technology emplacement began on April 16, 2001,
and was completed by April 25, 2001. Fieldwork
included Jackleg drilling and grouting 43 short
holes that were drilled in a radial pattern. A total
of 400 ft of drilling was completed during this
period, and 65 gallons of grout were injected into
these holes
Following the grout emplacement fieldwork, field
monitoring, field sampling, and data evaluation
continued through November 2003.
-------
MAGNETIC DECLINATION
14'SS EAST OF TRUE NORTH
I I I . * . I I
Figure 1-1. Miller Mine map.
-------
4*4+ 4 , *
|i?^^
** h
-f
4/It * * +,
t \ *
NTERMITTENT ELQWB
DISCHARGE
TO
GREENHORN CREEK
Figure 1-2. Generalized site map of the Miller Mine.
Figure 1-3. Photo of the Miller Mine upper and lower workings.
-------
Figure 1-4. Photo of the Miller Mine lower adit
discharge prior to implementation of the source control
technology.
Figure 1-5. Areas of AMD at the Miller Mine prior to the application of the
technology.
-------
2. Preinjection Site Characterization Summary
The primary purposes for characterizing the
underground mine workings were to provide
baseline information to determine where inflows
into the mine system occur, which structures
control the inflow, and how much inflow was
reduced as a result of the grout injection.
Geological, hydrogeological, geophysical, and
water quality information collected during the
mine characterization were used for evaluation of
the source control technologies, definitive design,
and field emplacement of the source control
material. The methods used to acquire both
background information and information for the
evaluation process are presented in the Activity
III, Project 8, Site Characterization and Materials
Testing Report (Ref 1), the QAPP (Ref. 2), and in
Sections 2 and 3 of this report.
2.1 General Information, Surveys, and
Observations
A topographic map and an underground mine map
were created from surveys completed at the Miller
Mine site. Mine mapping included completing
underground mine, geologic, and hydrogeologic
surveys and constructing maps of both the upper
and lower workings of the Miller Mine (Figure
2-1). A plan map showing the underground mine
workings, the continuous hydraulic monitoring
stations, and the main geological structure is
provided in Figure 2-1 and Appendix A, Plate 3.
2.2 General Site Geology
Tertiary igneous intrusive rocks (generally quartz
and granodiorites) intruded Precambrian
limestones and shales of the Newland Formation
(Belt rocks) to form Miller Mountain in the
Confederate Gulch and White Gulch area of
Broadwater County, Montana (Ref. 3). Structure
and topography of this area are the result of the
late Cretaceous and early Tertiary Laramide
orogeny that gave rise to folding, faulting, and
igneous intrusion. As a result, the dominant
structure affecting Precambrian rocks in the area is
the York Anticline (trend N35W). This structure
has been modified by successive reverse faulting
generally associated with the forced emplacement
of Tertiary igneous rocks that formed Miller
Mountain.
It is believed that the Miller Mountain intrusion
may have been emplaced in at least two stages
(Ref. 3). After the first period of intrusion, a steep
reverse fault developed that may be related to the
Miller Mountain Reverse Fault (Appendix A,
Plate 1). A second period of intrusion occurred
after the formation of the reverse fault that may
account for the secondary fracture pattern
described in this report. The secondary fracture-
shear pattern is likely related to later magmatic
injection into bedding planes and into preexisting
fault planes. This magmatic activity is likely
responsible for detachment, assimilation, and
shearing of Belt rocks into relatively small
remnants or residual masses of recrystallized Belt
rocks enveloped by intrusive rocks.
Mineralization in the Miller Mine underground
workings is associated with contact deposits along
or near the igneous-sedimentary contact.
Generally, the attitude of this contact within the
Miller Mine is N25W, 30SW.
2.3 Miller Mine Hydrogeology
Acid mine drainage discharge from the lower adit
likely originates from the Miller Mine Reverse
Fault (MMRF), associated fractures, and from the
secondary fracture-shear pattern related to
detachment and assimilation of Belt sediments.
These faults and associated fractures act as
conduits transporting groundwater through
overlying Precambrian Belt rocks and its
weathered equivalent into the underlying fractured
Tertiary quartz diorite (Tqd) (igneous intrusive
rocks) and lower mine workings. Results of
geologic mapping indicate that parts of the fault
plane have been injected with Tqd (Appendix A,
Plate 3). This likely occurred during the second
stage of intrusive activity during magmatic
injection of Tqd into bedding planes and the
associated detachment and assimilation of Belt
-------
rocks. Subsequent faulting and fracturing is
thought to have brecciated the Tqd and
Precambrian Belt rocks, creating fractures and
conduits through mineralized and altered rocks
rich in pyrite and other metallic sulfides.
2.3.1 Underground Mine Flow Regime
The lower workings of the Miller Mine is a 600-ft
haul tunnel designed for the removal of ore from
the upper workings and two attached drifts. The
drifts for this demonstration (Wl and W2) were
designated by the sample locations, which were
located at the mouths of the drifts (Figure 2-1).
The Wl drift is 68 ft long and was mined mainly
in the Tqd (Figure 2-1 and Appendix A, Plate 3).
The last 30 ft of the Wl drift runs parallel to the
Miller Mine fault system, and the flow was
measured using a small, 60-degree trapezoidal
flume manufactured by Plasti Fab, Inc.
The W2 weir was located at the mouth of the W2
drift and measured combined flows from the W3
weir as well as the flow from a winze that is
located between sample ports W2 and W3. This
winze is angled downward approximately minus
30 degrees to the north and was driven parallel
under the length of the northernmost section of the
W2 drift (Figure 2-1).
Pregrouting flow measurements show that only a
fraction of the total discharge measured
downstream near the adit portal at W4 is fromW2.
The average flow recorded at W2 was
approximately 1.15 gallons per minute (gpm)
(Appendix B). It should be noted that the flow
measurements at W4 could not be recorded during
the winter months and records the total discharge
from the Miller Mine adit portal, which is a
combination of W2 and Wl flows.
2.3.2 AMD Mine Discharge and Water
Quality
Metallic sulfide mineralization mapped in the
upper working level shows altered Belt rocks
replaced by pyrite and commonly altered to clay.
Rock core, logged during coring operations,
confirmed that sulfide mineralization and
alteration were common in Belt and Tqd rock
(Appendix C). As groundwater and oxygen
migrate through the mineralized fracture zones,
metals leach and acid forms, resulting in AMD
flowing into the lower underground mine
workings. As a result, elevated concentrations of
dissolved metals (Table 2-1) and reduced pH
levels are measured in the discharge water from
the Wl drift.
In contrast, Table 2-1 shows that the dissolved
metal concentration of water discharging through
the W2 weir, especially zinc (Zn), iron (Fe), nickel
(Ni), aluminum (Al), and manganese (Mn), were
several orders of magnitude less than waters
associated with the Wl drift. Using the
information from the underground geologic and
hydrogeologic surveys, it was determined that
water flows over calcareous Precambrian Belt
rocks (limey shale and siltstone) toward W3 and
W2 and remains neutral, resulting in the water
quality of flow from the W2 drift being below the
maximum contaminant level (MCL) for most
metals. As shown in Figure 2-1, it is not likely
that any water infiltrating the northernmost upper
workings comes in contact with mineralized rocks
associated with the igneous-sedimentary contact
zone; the mineralized rock is located below the
upper workings between core holes M6 and M4
(Appendix A, Plates 1 and 3).
Comparing the water quality prior to technology
emplacement between Wl and the adit discharge
measured at sample port W4 shows a decrease in
the concentration of dissolved metals (Zn, Mn, Fe)
and an increase in the pH measurements. The data
indicate that AMD and the concentration of
dissolved metals were reduced during the transit
between Wl and W4. Prior to implementation of
the technology, AMD exceeded the MCL criteria
for lead (Pb), and the secondary MCL for Fe at
W4.
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2.4 Geophysical Surveys
Geophysical surveys were conducted by DOE's
National Energy Technology Laboratory to
delineate water-filled fracture zones or surface-
connected mine voids that may underlie the survey
area at the Miller Mine. Initially, prior to
technology application, very low frequency (VLF)
and vertical-gradient magnetic surveys were
conducted as a reconnaissance effort to delineate
areas where surface water may be entering the
underground mine workings. Geophysical data
from the surveys was used to:
- identify fault/fracture zones or detectable mine
voids (i.e., stopes, drifts, etc.);
- provide a baseline geophysical signature that
could be used to evaluate subsequent
intervention activities (e.g., grouting);
- determine the direction of water flow in the
fracture along with the areas of high
mineralization;
- identify areas where additional work would be
needed; and
- identify the direction of groundwater flow
after the grout had been injected into the
underground workings.
The geophysical surveys performed prior to the
technology emplacement were conducted above
the underground gold mine using a 60- by
180-meter grid that was laid out on the steep slope
overlying the mine. This work produced
geophysical contour maps of the subsurface that
correlated the mine workings to the areas that had
the greatest potential for mineralization and water-
bearing fractures (Appendix D).
After the technology was applied, vertical-gradient
magnetic surveys were conducted, and data were
collected using a Geonic EM34-#XL conductivity
meter at intercoil spacings of 10, 20, 40, 50, 70,
and 80 meters using both horizontal and vertical
dipole orientations. The EM data was used to
model the apparent conductivity for the block of
earth underlying the surveyed grid. The modeling
was done using Emigma software in which the
conductivity images are configured as slices
suspended at specific depths below and following
the topography. Conductivity zones were
interpreted to indicate areas of fractured rock
where the increased conductivity was determined
to be from increased water content (AMD),
alteration (clay minerals that are conductive), or
sulfide mineralization (conductive sulfide
minerals) (Appendix D).
10
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MlaER MINE REVERSE FAULT
yWATER DRIPPING OUT OF RIB
NEAR CONTACT W/ Tkqd
MAGNETIC DECLINATION
14-55 EAST OF TRUE NORTH
X-CUTTING ALONG
CONTACT IRREGULAR;
SILUNC IN PLACES
BEDDING REPLACED W/PYRITE
>TO 1/2'; KAO ALT
tf Tkqd BELOW
-t-
EXTENSM
v.,-7 -WORKINGS
V" ABOVE
WATER WAS SOMETIME
I, m_ IN PAST DAMMED TO
NORTH AND RUN DOWN
WINZE (DRY 6/19/98)
WEIR OR FLUME LOCATION
W! - 60T TRAPEZOIDAL FLUME
W2 - 225" V-MOTCH WEIR
WJ - 225' V-NOTCH WEIR
W4 - 22.5' V-NOTCH WEIR
MILLER MINE RFVERSL" TAUIT
II
LI7
ACADf B9SHO/OB
RfVl D 6/S/W
DfWUR: H*S
Figure 2-1. Initial Miller Mine underground workings plan map showing the weir, flume, and sample port locations Wl,
W2, W3, and W4 (AutoCAD drawing B98R0708)
11
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Table 2-1. Background Water Quality for the Miller Mine Prior to Field Emplacement of the Source Control Technology
MWTP - Field Sampling Comparison
Site: Miller Mine
MWTP, Activity HI, Project 8
Sample Location: Comparison of All Sample Locations
Sample Date:
Sample #:
Sample Matrix:
Mine Inflow or Outflow
pH (su)*
Sulfate*
Dissolved Metals (mg/L, ppm)
Al
As
Cd
Cu
Fe
Pb
Mn
Ni
Zn
Ag
7/8/1999
Wl
Water
Inflow
6.08
828
0.145
0.025
0.003
0.002
81.600
0.023
5.970
0.097
0.915
0.004
7/8/1999
W2
Water
Inflow
6.87
295
0.012
0.025
0.003
0.002
1.330
0.023
0.594
0.010
0.057
0.004
7/8/1999
W3
Water
Inflow
7.50
445
0.010
0.025
0.003
0.002
0.014
0.023
0.072
0.010
0.013
0.004
7/8/1999
W4
Water
Outflow
6.58
539
0.010
0.025
0.003
0.002
18.600
0.023
3.220
0.046
0.388
0.004
MCL & Secondary
MCL, mg/L
6.5-8.5
250
0.050 - 0.200*
0.050 until 1/06
0.005
1.000*
0.300*
0.015
0.050**
0.088**
5.000*
0.100*
* Secondary MCL
** Aquatic Life Standard for fresh water
12
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3. Grout Injection Procedures
3.1 Introduction
Injection of a water-activated, expandable,
polyurethane grout into the fracture system in the
underground mine was completed in two different
stages, first in October 1999 and then in April
2001. The 1999 technology application included
core drilling and grouting 10 holes between the
W2 sampling port and Wl sampling port
(Appendix A, Plate 2). The April 2001
technology application complemented the first
injection and was performed to further reduce
AMD drainage from the Wl drift (i.e., drips,
weeps). All October 1999 work involved drilling
over and under the Wl drift. The April 2001 work
involved Jackleg drilling and grouting holes
drilled directly in the Wl drift adjacent and into
the MMRF. The following sections describe core
and rock drilling core recovery, water injection
tests, grout formulations, the grout injection
system, the grout injection process, and the
monitoring performed to observe the affects of
grout injection. A discussion of the performance
of the test, the water and dye injection
observations, and the grout injection are presented
in Section 4.
3.2 October 1999 Core Drilling and
Recovery
Drilling and coring was accomplished using a
Hagby 1000 core drill and Gardner-Denver Model
83 Jackleg drill. Rock core from nine core holes
was recovered and described (Ml, M2, M3, M4,
M5, M6, M7, M8, M9) (Appendix A, Plate 2). An
additional hole was drilled using the Jackleg drill
(MJ10). Rock core was not recovered from MJ10.
Approximate drill locations were marked along the
eastern rock face between mine survey points LI 1
and L12 (core-drill locations) and between LI 1
and L13 (MJ10) (Appendix A, Plate 2).
Respective bearings of each core hole were
determined by orienting the longitudinal axis of
the Hagby core drill parallel to a string stretched
between two spads attached on the eastern and
western faces. The compass bearing between
respective spads was determined using a Brunton
compass prior to the setup and arrival of the core
drill and associated equipment.
Setup for each core hole included fastening the
front frame of the drill rig to the rock face using
two Dywidag rock bolts (approximately 1-inch-
diameter threaded steel) embedded approximately
4 ft deep into the rock face with Fastloc epoxy.
This setup provided stability for the core drill and
ensured that coring operations would be successful
and safe.
Water circulation was established and used as the
drilling and coring fluid. Water was pumped from
the flooded winze located north of the W2 weir
into a plastic reservoir and then drawn using a
hydraulically powered reciprocating pump and
injected through the tools of the drill string and
bit. Fluid pressure and flow rate were regulated on
the Hagby control panel to achieve the desired
flow circulation, feed rate of bit, bit pressure, and
rotation rates of the drill rods and diamond bit.
These parameters, depending on the size and type
of drill tools, were optimized for the best core
recovery. Drill cuttings were flushed from each
hole by spent drilling fluids.
3.2.1 Recovery of Orientated Core
Continuous and orientated core was recovered
from three core holes (Ml, M2, M3) using the
clay-impression method and associated tools
(Appendix C). Orientated core was recovered
using a clay core orientor tool constructed from
the inner tube of a core barrel. This tool is
weighted with Pb the full length of one side of the
barrel, which forces the weighted side down.
Orientation was accomplished by placing potter's
clay on the shoe of the orientation tool, inserting it
in the hole, and then forcing the clay to make an
imprint of the base of the hole. This impression
was used to align the orientation of the next core
sample taken with that of the last core run.
Recovered core was subsequently laid out in a
split metal tray and its "up" orientation was
13
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scribed the full length of the core run.
Occasionally, it was not possible to orientate or
scribe the entire core run due to intervals that were
severely fractured, altered, or weathered.
3.3 Water Injection and Dye Tests
Water injection tests were performed in four core
holes (Ml, M2, M3, and M8), and dye tests were
performed in two core holes (M3 and M4) using a
blue food-grade dye. These tests were done for
design purposes to determine the relative hydraulic
conductivity and the interconnection of the
fracture system (Appendix A, Plate 4).
Generally, water was pumped into a completed
core hole through a packer system configured for
the desired test interval(s) for water injection
testing (Ref 4). Pressure and flow rate were
regulated to achieve the desired test parameters for
each unique testing interval. Water used for
testing was pumped from the winze located north
of Weir 2 into a plastic reservoir tank. Water was
drawn from the reservoir tank with a hydraulically
powered pump and transferred into the respective
core holes through the packers (behind or
between) and injected through 1/2-inch inside
diameter (ID) perforated steel pipe (perforations
on 6-inch centers) (Figure 3-1). Pressure was
regulated at the Bean 20 pump with a built-in
pressure control valve. A hydraulic motor control
on the Hagby control panel regulated the flow rate
of the reciprocating pump for water testing, and a
Great Plains Industries, Inc. electronic digital
meter was used to record total water volume per
test. A glycerin-filled pressure gauge was used to
record test pressure.
3.4 Grout Formulations and Emplacement
Grout was injected into core holes as specified
through a grout delivery system that consisted of a
plastic reservoir tank "fill-rite" flowmeter
connected in-line to the reservoir outlet (suction
side of pump), and a hi-con dual-action piston
pump (Figure 3-2). The grout delivery system was
used to inject grout into the respective core holes
through the packer assemblies and at the desired
interval. An additional pumping system was used
to supply and mix water (necessary for grout
activation) in tandem with grout emplacement at
the core hole standpipe.
The water delivery system consisted of a sump
pump placed in the winze located north of weir 2,
which conveyed water into a plastic tank. Water
was drawn from this tank with an air-powered
Bean pump and plumbed to a tee located at the
core hole standpipe. Grout was injected through
the run of the tee with water injected at the branch
of the tee. Generally, grout and water were
injected simultaneously into the core hole; the
grout delivery valve was open, and water (supplied
by the additional pumping system) was injected
into the core hole. The "fill-rite" flowmeter was
used to record total grout volume injected into the
core holes. A pressure gauge located at the
standpipe was used to record injection pressures of
both water and grout. The Great Plains Industries
flow totalizer recorded the volume of water added
to the core hole.
3.5 April 2001 - Wl Drift Drilling and
Grouting
A Gardner-Denver Model 83 Jackleg drill was
used to drill the radial pattern of grout injection
holes used to seal the fracture system during the
April 2001 technology emplacement.
Approximately 400 ft of hole were drilled and
45 gallons of grout were injected into the Miller
Mine fracture system at pressures between 60 and
300 pounds per square inch (psi). The objective
was to determine if the small seeps that existed
after the initial grout injection would be
minimized and the flow from the Miller Mine adit
effectively reduced further. The Jackleg drill was
used because it would fit into the small Wl drift
(4-ft by 6-ft drift). Mechanical packers were used
instead of pneumatic inflatable packers because of
the ease of use and cost effectiveness.
14
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3.6 Geophysical Investigation Methods
The geophysical techniques used for this
demonstration project included VLF terrain
conductivity and field gradient magnetometry.
The field gradient magnetometry was performed in
the underground mine workings of the Miller
Mine. The survey was performed in the upper and
lower mine workings. The gradient magnetometry
was used to distinguish anomalies due to
mineralized zones, which the gradient
magnetometry can detect, from anomalies due to
water-filled fractures, which gradient
magnetometry cannot detect. The VLF is sensitive
to conductive areas at depth. These areas include
anomalies due to mineralization and water-filled
fractures. This combination of techniques
provided a quick, comprehensive approach for
delineating conductive zones.
Figure 3-1. The drilling/grouting subcontractor placing an inflatable packer
system into the grout hole for dye and water injection testing.
Figure 3-2. Photograph of the grout pump, tanks, and flowmeter system for
the Stage I grout injection in 1999.
15
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4. Phase I - Field Emplacement Results
Field emplacement and associated fieldwork was
performed by MSB and Bush Drilling
Incorporated in two phases. Phase I fieldwork
consisted of drilling, coring, logging, testing, and
grouting 10 boreholes placed in a radial pattern at
select locations in the lower working level.
Support equipment included a generator,
compressor, core and Jackleg drills, mucker,
various pumps, tanks, lines, ancillary tools,
equipment, and supplies.
The HA Combi grout, manufactured by De Neef
Construction Chemicals, Inc., was used for the
demonstration. The HA Combi grout is available
in 5-gallon, closed-head pails sealed under dry
nitrogen since the grout is a water-activated
material. The HA Catalyst Accelerator was used
at 2% of grout volume. The volume of catalyst
can vary depending on the designed activation
time allotted for the project. For this project, 2%
catalyst was used.
Bush Drilling provided all personnel (driller and
helper), equipment, and tools necessary to drill,
test, recover rock core, and perform grouting
operations. MSB site personnel included a field
team leader/site safety officer, site security, and
technical personnel.
4.1 Preliminary Design (1999)
In Phase I, Technology Deployment (1999),
drilling, coring, water and dye testing, and
grouting of the associated fracture systems in the
lower working level of the Miller Mine was
accomplished in three design segments and at
three drill stations (Appendix A, Plate 1). Design
Segment 1 included drilling, coring, water testing,
and grouting Ml through M4 at the first drill
station located between survey points LI 1 and L12
(Appendix A, Plate 1). Design Segment 2
included drilling, coring, and grouting M5, M6,
and M7 from the first drill station. Design
Segment 3 included drilling, coring, and grouting
M8 and M9 from a second drill station located just
east of survey point LI 1, and MJ10 was drilled at
a third drill station located approximately 33 ft
northeast of weir 1 in the Wl drift (Appendix A,
Plate 2).
The respective core holes were drilled to intercept
the fracture system associated with the Wl drift
and the MMRF. The core holes were drilled
perpendicular to the fracture system. Only three
drill stations were used to minimize mobilization
and movement of the drill rig. The fracture system
intersected by the core holes was mapped in the
Wl drift during the initial underground survey
prior to drilling.
Typically, the work sequence for each of the nine
core holes included drilling and recovering core,
descriptive logging, performing water injection
and dye tests in three core holes, and grouting.
Core was not recovered from the last borehole
(MJ10) since Jackleg drilling methods did not
allow for the recovery of core. Approximately
624 ft of drilling was completed, which included
617 ft of coring. Core recovery ranged between
94% and 100% for each of the nine core holes
(Table 4-1).
The primary goal of Design Segment 1 work was
to shut off or significantly reduce fracture flow of
AMD into the Wl drift and make work associated
with Design Segment 2 unnecessary. However,
after grouting all the grout holes at Design
Segment 1, the results indicated that the flow in
the Wl drift had only a 75% reduction when
averaged over al0-month period and compared to
the original flow volume. As a result, Design
Segment 2 drilling and grouting was performed to
further reduce flow from the Wl drift in an
attempt to achieve the 95% reduction criteria.
While performing the Design Segment 1 work, the
local direction of fracture flow into the Wl drift
was identified and mapped. During Design
Segment 2, it was determined that the main inflow
into the Wl drift occurred on the right-hand rib or
eastern side of the drift, and therefore Design
Segment 2 grout injection was performed to
16
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reduce the inflow on the eastern side of the Wl
drift. To further reduce the inflow from the east
side of the Miller Mine Wl drift, Design Segment
3 work was undertaken and entailed performing
Jackleg drilling and grouting drill hole number
MJ10.
4.1.1 Design Segment 1 - Grout Injection
Design Segment 1 work consisted of setting up the
core drill at Drill Station No. 1 and drilling a radial
pattern of four core holes (M1-M4) from four
pivot points all located within 5 ft of each other.
These initial core holes radiated from N53E to
S81E and were inclined up from the horizontal (9
degrees to 13 degrees). Each borehole was capped
with standpipes, cored, pressure tested with water,
and grouted as described below. This primary
phase of drilling and coring included a total depth
of nearly 296 ft (Table 4-1 and Appendix C)
Design Segment 1 grouting initially reduced the
average fracture flow discharge from 5.88 gpm to
1.24 gpm as measured at the Wl sample port
(Appendix B). In addition, data from this work
indicated that an increased flow occurred in the
southeastern section of the area to be grouted
(Appendix A, Plate 4). This became evident after
Ml, M2, and M3 were drilled and cored.
Discharge progressively increased from zero in
Ml (northernmost core hole) to approximately
4 gpm in M3 located south of Ml.
4.1.2 Design Segment 2 - Grout Injection
Design Segment 2 grout injection work consisted
of drilling three core holes (M5, M6, M7) from the
first drill station. These core holes radiate from
due east to N85E and were completed to reduce
the flow beyond that observed from Design
Segment 1. However, significant reductions of
AMD discharge were not observed after
completion of the work. Influx into the Wl drift
varied between 1.53 gpm and 1.03 gpm (0.10 and
0.08 Wl vertical head measurement) directly
following completion of grouting at M5, M6, and
M7.
4.1.3 Design Segment 3 Grout Injection
Design Segment 3 grout injection was performed
to reduce the flow entering the right rib (east side)
of the Wl drift. The work consisted of drilling
two core holes from the second drill station
(located just east of survey point L11; grout core
holes M8 and M9) nearly parallel to the Wl drift
but with different inclinations. Additionally, one
short hole (MJ10), located approximately 33 ft
northeast of the Wl flume, was drilled and
grouted.
4.2 Rock-Core Drilling and Recovery
The Hagby core drill was used to recover rock
core from nine core holes (Ml, M2, M3, M4, M5,
M6, M7, M8, and M9) located on the east rib of
W2 between mine survey points LI 1 and L12
(Appendix A, Plate 1) within the lower workings
of the Miller Mine. An additional borehole
(MJ10) was drilled with the Jackleg drill in the
east rib of the Wl tunnel approximately 33 ft
northeast of the Wl weir.
In general, the rock core was described as a
fractured and altered chloritized quartz diorite or
granodiorite (Miller Mountain Intrusive rocks)
with minor amounts of fractured silicified
claystones and mudstones (Precambrian Belt
rocks). Core recovery and details about each hole
are shown in Table 4-1, and core logs are included
in Appendix C.
Two sizes of core were recovered from each core
hole (HQ-NQ3 or BQ-AW34); slightly larger core
(HQ, BQ) was recovered the first 10 ft of each
core hole (Table 4-1). The larger size core (HQ or
BQ) was drilled and recovered in order to install a
10-ft standpipe with a threaded collar at the face.
Core holes Ml, M2, and M3 had HQ size
standpipe installed [3.764-inch outside diameter
(OD)]; M4 through M9 had BTW standpipe
installed (2.346-inch OD).
All core was logged, photographed, and stored in
water-resistant core boxes. Core logs included
lithologic descriptions of the rock core and
drawings of the respective fracture patterns
17
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(Appendix C). Orientated core was recovered in
the first three core holes (Ml, M2, M3) (Table
4-2) using the clay-impression method. Clay
impressions were taken at the base of the core hole
after each core run, matched to recovered core,
and used to determine the orientation of the core
and the fractures that were to be grouted.
4.2.1 Downhole Camera Survey
A Sperry-Sun single-shot borehole deviation tool
was used to confirm the bearing and inclination of
Ml, M2, and M3 completed during Design
Segment 1. This tool used a downhole camera to
provide a photograph at discrete depths of the
magnetic orientation (compass bearing) and
inclination of the borehole. Generally, data from
this borehole deviation tool confirmed the initial
bearing and inclination determined by the Brunton
compass. This tool provided the compass bearing
relative to magnetic north. A correction factor of
17 degrees was added to correct for the local
magnetic declination of 17 degrees east of true
north. Table 4-2 details the results of the
downhole survey.
4.2.2 Core Hole Location Survey
A theodolite was used to determine the location of
each initial drill point upon the rock face above the
sill (northing, easting, and elevation). This
information is shown in Table 4-3. Generally, the
initial drill point for Ml through M6 was
approximately 5 ft above the sill. The initial drill
point for M7 was approximately 3 ft above the sill
and over 6 ft for M8 and M9.
4.2.3 Geology - Recovered Rock Core
Approximately 617 ft of rock-core (sedimentary
rocks of Precambrian age and Tertiary igneous
rocks) were recovered, described, and logged
during drilling and coring operations. The
dominant rock type included a quartz or
granodiorite likely emplaced during the Tertiary
period by forceful injection into Precambrian
limey shales and mudstones. Many of the rock
fractures logged and described in the rock core are
likely related to intrusive emplacement and
associated faulting and deformation.
Generally, the granodiorite was slightly to
extremely altered, weathered, and had a slight
greenish cast due to chloritization. Previous work
confirms that hydrothermal alteration of biotite
and hornblende to chlorite is common (Ref 3).
Numerous, but generally small intervals of core,
were extremely altered, soft, weak, and clayey.
This is due to the complete alteration of orthoclase
to kaolinite and plagioclase to sericite (Ref. 3).
Many of these altered intervals included sheared
fracture planes with slickenside textures that often
occurred in the vicinity of the MMRF. Generally,
these rocks are extremely fractured and cut by
numerous quartz veins and veinlets that are
commonly mineralized with Fe and Cu sulfides.
Disseminated sulfide mineralization was common
throughout the core.
Silicified and fractured shales, mudstones, and
claystones were logged and described in the rock
core. These rocks comprised approximately 52 ft
of the total potential core of 617 ft and generally
were creamy white to creamy pale brown, hard,
and suffused by numerous quartz veins and
veinlets to give a contorted banded appearance.
These rocks are mapped in Appendix A, Plate 3,
which shows in plan view intrusive and silicified
sedimentary rocks and associated faults and
fractures. Appendix A, Plate 3 also shows, based
on rock core data, the relationship between Belt
and igneous intrusive rocks. Generally, Plate 3
supports the concept of pressurized magmatic
injection along zones of weakness as well as local
magmatic assimilation of Belt sedimentary rocks
during periods of intrusion. Remnants of silicified
Belt rocks are enclosed or otherwise surrounded
by igneous intrusive rocks. Fracturing and
faulting are concentrated in areas associated with
the contact between the sedimentary and igneous
rocks.
4.2.4 Structure
Detailed geologic and structural interpretation
within the Miller Mine is based on the work of Mr.
E. A. Johnson (Ref. 3) and recent work by MSB
during coring and grouting operations within the
18
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mine. Recent structural data was gathered from
the nine radiating core holes between mine survey
points LI 1 and L12. In general, these core holes
were drilled normal to the main structural trend
associated with the MMRF having a trend
approximately N30W with a dip approximately
80 degrees west. Numerous fractures and faults
are associated with the contacts between the
igneous and sedimentary rocks (Figure 2-1 and
Appendix A, Plate 3).
4.2.5 Fracture/Shear System - Mine Map
Fractures in the granodiorite mapped in the Wl
drift dip steeply toward the west and show two
major strike patterns of N15W to N60W and
N65W to N40E. Generally, fractures mapped in
the Wl drift are located 2 to 7 ft apart and become
more abundant and closely spaced toward the
northeast (near the MMRF). Fracture origin is
likely due to movement along faults and the
development of stresses within the zone of the
MMRF.
4.2.6 Fracture/Shear Systems Rock Core
Plate 3 of Appendix A, which was developed from
core data, shows two fracture systems associated
with the MMRF. The first of these systems has a
strike pattern of N10W to N45W while the second
has a strike pattern of N30E to N45E. Core data
indicated that both the granodiorite and silicified
sedimentary Belt rocks are highly fractured and
slightly to extremely altered by the actions of hot
water
Plate 3 of Appendix A, shows several areas of
undifferentiated Precambrian Belt rocks enclosed
by Tqd and generally bordered on the east and
west by fractures. It is believed that this fracture
pattern may be one of the primary water conduits
into the Wl drift from the MMRF. Several
fractures shown on Plate 3 intersect the Wl drift
and are associated with water-bearing intervals.
The presence of this group of fractures could not
be determined from visual observations or
mapping conducted in the Wl drift.
Average fracture density described from core
samples varies between two to nearly seven
fractures per foot of core. This is much greater
than the fracture density mapped in the Wl drift
and is possibly due to the masking effects of the
abundant coating of Fe oxides and other alteration
products along the drift. Generally, fracture
density associated with core samples increased
with drill hole depth and in the direction of the
MMRF (Figure 2-1 and Appendix A, Plate 3).
These closely spaced fractures are generally steep,
dipping between 60 degrees to near vertical,
especially near the MMRF. Core sections near the
MMRF are extremely soft, weak, altered, and
clayey. By contrast, core sections taken at a
distance from the MMRF, but associated with the
secondary fractures, are generally fragmented with
a striated texture (slickensides) and not altered to
clay.
4.2.7 Observed Water-Bearing Fractures
Water-bearing intervals were recorded while
drilling and coring in the lower workings of the
Miller Mine. Details are included in Table 4-4 and
Appendix A, Plate 4. Results of core drilling
indicated that water-bearing fractures occur
southeast of Ml, and flow increased from virtually
zero in Ml to over 4 gpm in M3 and M4 near the
zone of the MMRF (Plate 4). Water flow was
observed from all core holes except for Ml and
M5 (Table 4-4, Table 4-5, and Appendix A, Plate
4). It is likely that a significant part of this water
is associated with water-bearing fractures
associated with the MMRF and possibly the
contact between igneous and silicified Belt
country rocks. Core hole data indicated that local
water flow within the fractured rocks generally
increases to the southeast.
Water-bearing intervals of fractured rock, recorded
from field data, are presented in Table
4-4 and mapped in Appendix A, Plate 4.
Generally, all recovered core was fractured and
slightly to extremely altered. Water-bearing
fracture systems generally were found to occur
beyond 40 ft within core holes M2, M3, M4, M6,
M7, M8, and M9 in the vicinity of the MMRF.
Appendix A, Plate 4 shows water-bearing
19
-------
intervals, the mapped location of each core hole
and potential pathways between core holes.
4.2.7.1 Description - Water-Bearing Fractures
Water-bearing fractures occurred in seven core
holes (M2, M3, M4, M6, M7, M8, and M9) and
are described below. Except in M4 and M9,
water-bearing fractures were found to occur in the
quartz-granodiorite. Many fractures associated
with M4 and M9 are related to the diorite-Belt
rock contact. Grout hole M5 was primarily a dry
drill hole; however, fractures in M5 were
connected to the fractures in M4 and M6. Cross-
communication between M5, M6, and M4
occurred during grout injection.
4.2.7.2 M2 Water-Bearing Fractures (44 to
49ft)
Fractures in this bore hole interval are the result of
shear associated with faulting (slickenside texture-
surface smooth, glassy finish with parallel
striations) and are narrow, open, and slightly
coated or stained with Fe oxide, chlorite, or clay
(kaolinite). Much of the core from this interval
can be described as a very altered diorite with a
closely spaced fracture pattern that is highly
broken and fragmented (Log Description of M2,
Appendix C).
4.2.7.3 M3 Water-Bearing Fractures (53 to
63 ft and 68 to 73 ft)
Rock fractures in the 53- to 63-ft interval of this
bore hole occur in chloritized diorite. Fractures
are partly the result of shear (described above) and
are narrow to moderately wide (0.05 to 0.5 inch in
width), unstained to partially filled with Fe oxide,
Mn oxide, clay, chlorite, and quartz. These
fractures are open and permeable.
Rock core from the 68- to 73-ft interval is sheared
with abundant slickensides, soft, weak, and
extremely altered to clay. Fractures are often
partially filled with chlorite, clay, and quartz.
4.2.7.4 M4 Water-Bearing Fractures (23 to
31 ft and 45.5 to 46.5 ft)
Rock fractures in the 23- to 31-ft interval of this
bore hole include a sequence of fractured silicious
Belt rocks (M4 Log, Appendix C). Water-
producing fractures in this interval area are
believed to be associated with the diorite-Belt rock
contact. Typically, fractures near the contact are
closely spaced and have produced fragmented
rubble, especially between 24 and 25 ft. These
fractures are narrow to moderately wide and
stained and partially infilled with Fe oxides and
clay. The entire interval has been shot through or
suffused with quartz veins to 0.5 inch in thickness
and mineralized by disseminated Cu and Fe
sulfides.
Water-bearing fractures in the 45.5- to 46.5-ft
interval of this bore hole are associated with
fractured diorite, narrow with a striated texture
(sheared - slickensides), and slightly coated with
clay, chlorite, or manganese oxide.
4.2.7.5 M6 Water-Bearing Fractures (43 to
44ft)
Fractures associated with this bore hole interval
have fragmented the core and are stained with
oxides of Fe.
4.2.7.6 M7 Water-Bearing Fractures (52 to
57ft)
Fractures associated with this bore hole interval
are similar to fractures described above in M6,
except that these fractures are stained or partially
filled with clay and manganese oxides. As
described above, the core is chloritized, suffused
with quartz veins to 2 inches in thickness, and
mineralized with disseminated sulfides of Fe and
Cu.
20
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4.2.7.7 M8 Water-Bearing Fractures (48 to
53ft)
Fractures associated with this bore hole interval
are associated with shear, especially near 49 ft.
These fractures are generally narrow and stained
to partially infilled with clay and chlorite. This
interval is chiefly chloritized diorite as described
above, but includes a small interval, between 51
and 52 ft, of silicified Belt rocks.
4.2.7.8 M9 Water-Bearing Fractures (31 to
32 ft and 43 to 58 ft)
Fractures within the 31- to 32-ft interval of this
bore hole are associated with the contact between
altered diorite and silicified Belt rocks. This
interval is rubblized with fractures that are narrow,
open, and stained with oxides of Fe.
Fractures within the 43- to 58-ft bore hole interval
are narrow to moderately wide and stained to
partially filled with oxides of Fe, chlorite, and
kaolinite. Much of the core within this interval is
extremely altered, soft, and clayey ; water-bearing
fractures between these extremely altered sections
are rubblized and completely fragmented.
4.3 Water/Dye Injection Tests
Water/dye injection and testing of select intervals
(stages) was done in several core holes before
grout injection (Ml, M2, M3, and M8) using
packers inflated with nitrogen to pressures of
approximately 150 psi to 200 psi. Water/dye
injection tests were conducted to:
- determine the approximate hydraulic
conductivity of the fractured rock;
- confirm that specific intervals and logged
fracture patterns were interconnected between
core holes;
- confirm that fractures were able to take water;
and
- confirm that fractures had the potential to
accept grout.
Water mixed with food-grade blue dye was
injected into M3 and M4 as part of the water
injection testing phase. Blue dye injection testing
was used to:
- more easily see any cross communication or
interconnection between core holes;
- determine where water entered into the Wl
drift;
- determine the approximate travel times from
injection points to discharge points; and
- determine the approximate volume of
discharge from respective discharge points.
4.3.1 Lugeon Water Injection Testing
The Lugeon method was used to calculate the
relative permeability of the fractured interval
during water injection testing (Ref 5). This
method provides information on whether or not a
fractured interval will easily take water based on
the amount of water injected per unit length of the
interval per unit time (liter/meter/minute).
Generally, a value of 1 Lugeon is a low
permeability area where grouting is not necessary;
10 Lugeons warrant grouting for most seepage
reduction jobs. A Lugeon value of 100 indicates
areas where fractures are common and open and
grouting is necessary to control seepage.
The water testing Lugeon values and borehole
permeabilities are presented in Appendix E.
Lugeon values shown in Figures 4-1, 4-2, and 4-3
were calculated as a function of injection pressure:
the Lugeon value is 2 (liters per meter per 5
minutes) when injection pressure was 1 bar (15
psi); when pressure was other than 1 bar, the value
was liters/meter/minute)(10)/actual pressure in
bars.
High Lugeon values indicate relatively high
fracture permeability for a given interval. For
example, Figure 4-1 shows a comparison of
Lugeon values and the total number of fractures of
several cored intervals in bore hole Ml. The
interval between 40 ft and 70 ft shows a relatively
low Lugeon value (8) associated with a high
fracture density (6 fractures per foot).
21
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By contrast, the interval near the face of Ml (10 to
18 ft) has a high Lugeon value (greater than 100)
associated with a low fracture density (2 fractures
per foot). This data illustrates that fractures in the
40- to 70-ft interval are relatively tight and
impermeable compared to fractures associated
with the 10- to 18-ft interval. Overall, the Lugeon
value of bore hole Ml decreases with hole depth
as the fracture density increases.
The Lugeon value of bore hole M2 is relatively
constant through the first 52 ft of core (61-66) and
then drops significantly between 52 and 70 ft of
bore hole to a value of approximately 12. The
fracture density is similar throughout the bore hole
with an increase from 4.2 to 5.5 in the 32- to 52-ft
section. Figure 4-3 shows that the Lugeon value
for bore hole M3 drops from greater than 100 to an
average of approximately 50 after the first 20 ft of
bore hole. The fracture density for this hole is
nearly constant through the first 68 ft at a value
near 4.5. The fracture density increases in the last
section of the bore hole to a value of 6.8.
Generally, Figures 4-1, 4-2, and 4-3 indicate a
decrease of fracture permeability with depth
(decreasing Lugeon values) associated with a
slight increase of fracture density. This may
indicate that an open and more permeable fracture
pattern was developed near the rock face during
blasting and mucking associated with the
excavation of the mine working.
4.3.2 Lugeon Fracture Hydraulic
Conductivity Results
Fracture hydraulic conductivity values were
calculated based on the Lugeon values calculated
from water-injection tests. Lugeon hydraulic
conductivity values calculated for the fracture
pattern associated with core holes Ml, M2, M3,
M4, and M8 ranged between 10"3 and 10"4
centimeters per second (cm/s). These values were
calculated based on the approximate relationship
of 1 Lugeon unit equals 1.3 x 10"5 cm/s (Ref 5).
Specific values are listed in Table 4-5.
Generally, these values indicate that the
conductivity of the fracture system using the
Lugeon method is similar to the conductivity of
sandy silt. However, the hydraulic conductivity of
the system is also a function of a fluid's density
and viscosity. Lower hydraulic conductivities of
the fracture pattern may be expressed as density
and viscosity of a given fluid increase. This is the
case since the grout selected for injection has a
viscosity (500 to 700 centipoises (cps) at 77 °F),
which is two orders of magnitude greater than the
viscosity of the mine waters (1.4 cps at 41 °F)
used during the injection tests. Due to the high
viscosity of the grout, the hydraulic conductivity
value for the fracture system is reduced, thus
limiting the effective radius of the grouted area
(Table 4-6). This is especially true since the great
majority of fractures logged and described were
narrow and tight (Appendix C), and very little
grout was observed in fractures within the core
from previously grouted holes as discussed below
and shown in Appendix A, Plate 4.
Preliminary calculations indicate that the hydraulic
conductivity of the fracture system was
approximately two to three orders of magnitude
lower when a viscous grout was substituted for
water. For example, all hydraulic conductivity
values noted in Table 4-5 are reduced to
approximately 10"6 cm/s when calculating the
hydraulic conductivity of the fracture system using
a viscous grout, thus the radial extent of the grout
emplacement is reduced.
4.4 Phase I - Grout Emplacement
Grout injected into each core hole (M1-M9)
consisted of an FŁA Combi polyurethane grout
combined with a FŁA Flex Cat Accelerator
(fast/slow activator) that was designed to fill voids
and fractures. This material activates and expands
on contact when mixed with water. Field
observations of this material indicated that this
material had a common expansion ratio of at least
5 to 1 when thoroughly mixed with water.
This grout was chosen based on its performance
during laboratory tests. Criteria critical to the use
of this material was its ability to withstand and
solidify under cool temperatures known to occur
22
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in the Miller Mine and when submerged in acidic
waters. This grout passed all test requirements and
was selected from among 50 other grouts
subjected to identical tests.
The HA Combi Grout is a resin activated by a
urethane activator (HA Flex Cat); lines, hoses,
tanks, grout pumps, and sundry equipment were
cleaned with a special washing agent after each
grouting operation. This washing agent is a
nonflammable solvent mixture necessary to clean
the injection equipment and prevent plugging by
residual activated grout.
All grout injected into each respective core hole
was activated by the fast urethane activator except
for the 32- to 70.5-ft interval in M2. Each batch of
grout was mixed with a 2% portion of activator.
For example, each 5-gallon batch of grout was
mixed with 0.10 gallons of fast activator or
0.38 liters. Table 4-7 details the amount of grout
injected into each core hole, amount of water
injected with the grout, average injection pressure,
time required to inject, and date of injection.
A grout sample was recovered from each grout
batch prior to injection. These samples, each
approximately 3 ounces, were thoroughly mixed
with a similar amount of water to determine
reaction times. Generally, grout reaction and
associated expansion began 2 to 3 minutes after
being mixed with water between 65 °F and 70 °F.
Observed grout expansion under these conditions
ranged between 4 and 5 times the original volume.
Table 4-6 gives general properties of the grout,
and Table 4-7 reports grout injection data
associated with each core hole.
This data includes the injection intervals, amount
of grout and water injected, and average injection
pressure. Using Table 4-7, the sequence of grout
injection for each core hole between October 4 and
October 20, 1999, can be followed.
4.4.1 Core Hole Fracture Cross
Communication
Water testing and later grout injection confirmed
that the fracture system, described and logged at
each core hole, would accept grout, and
preferential communication existed between core
holes. The water and grout injected into each
respective core hole followed a fracture path of
least resistance, and many of the fractures
described in the core were too narrow and tight to
allow the viscous grout to fill and seal the fracture
system. Fracture communication and fluid
transport were observed during the water injection
testing and the grout emplacement. The
communication between fractures is greatly
influenced by the viscosity of the fluid (water or
grout) and perhaps by capillary forces associated
with tight and narrow fractures. Cross
communication between core holes was faster and
more widespread as the viscosity of the fluid was
reduced. For example, communication between
core holes or discharge into the Wl drift was
much less when a viscous grout was substituted
for water during grout injection. Fracture
communication and interconnection between core
holes was inferred while drilling and coring, water
testing, and during grouting. Appendix A, Plate 4
shows details of fracture communication recorded
in the field.
Fractures observed in the core were logged and
described in detail. With few exceptions, fractures
were narrow (0.05- to 0.1-inches in width) to very
narrow (less than 0.05 inch in width). Fractures
occurred in all recovered core. Generally, these
fractures were narrow, stained with Fe, and not in-
filled. Maximum fracture density ranged from
over four fractures per foot to nearly seven
fractures per foot in the first three core holes (Ml,
M2, M3) (Figures 4-1, 4-2, and
4-3). These fracture densities are representative
and similar to fractures logged in M4, M5, M6,
M7, M8 and M9 (Appendix C).
4.4.1.1 Fracture Communication - While
Drilling and Coring
Fracture communication between core holes was
observed between Ml and M2 while drilling M2.
In general, water flow (chiefly from drilling fluid)
increased slightly in all core holes and from the
Wl drift while drilling and coring M2-M7
(Tables 4-4 and 4-8). Appendix A, Plate 4 shows
23
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the direction, distance, and packer intervals where
communication of grout or water occurred during
application of the technology or while drilling
grout holes.
Congealed grout was found as a fracture filling in
five different, very small intervals among four
closely spaced core holes (Appendix A, Plate 4).
However, this grout was gelatinous and did not
resemble or have the texture of grout samples left
to congeal in a nonconfmed environment. This
indicates that much of the grout did not travel
through the tight and narrow fracture pattern but
took a preferred path. The single exception was
the grout injected into MJ10 (drilled with a
Jackleg drill). This grout sealed a nearby weep
hole having an inflow of 1 to 2 gpm in the Wl
drift immediately after injection and had the same
texture as grout left to congeal in a nonconfmed
environment. Apparently, fractures associated
with this weep hole were relatively large, open,
and allowed for grout transport.
Grout, likely from M3, was observed in small
intervals of core from M4 and M5 (Appendix A,
Plate 4). Similarly, grout was described as a
fracture filling in a small interval of core from M9
that likely originated from M8 or M6 (Plate 4).
Likewise, two small intervals of grout were
described as a fracture filling in M6 that likely
originated from M5 (Plate 4).
An aqueous discharge of approximately 4 gpm
was observed from grout hole M3 immediately
after completion drilling. This observation
correlated to a reduction of discharge from the Wl
weir and indicates that there is a direct relationship
between flow measured in the Wl weir and
fractures intercepted by M3.
Similarly, an increase in the discharge to the Wl
drift was observed immediately after M4 was
cored. This was a direct result of M4 skimming
and breaking through the back of the Wl drift roof
and intercepting a water-producing fracture that
previously had very little affect upon
measurements of the Wl weir. This resulted in an
increase in the discharge from the Wl weir as
water (approximately 4 gpm) rained from the back
where M4 had broken through the roof of the mine
workings. This affect was transitory; discharge
from the Wl weir was reduced after M4 was
grouted.
4.4.1.2 Fracture Communication While Water
Testing
Water testing was performed on Ml, M2, M3, M4,
and M8. This included injecting blue dye (food
coloring) mixed with water into M4 and several
intervals of M3 (Table 4-8). Fracture
communication between core holes was observed
between M2 and M3 during water injection testing
of M3 (Table 4-8). Increased water flow was
observed from M2 while testing the 40- to 70-ft
Ml interval. Similarly, increased water flow was
observed from M2 while testing several shallow
intervals of M3, and blue dye from Ml and M2
was observed while testing deeper intervals of M3
(Table 4-8 and Appendix A, Plate 4). No increase
in water flow was observed from M3 or Ml while
testing M2. Also, because there was
communication between M3 and Ml grout holes,
it was apparent that there was a large flow path in
the deeper intervals setting close to the MMRF.
4.4.1.3 Fracture Communication - While
Grouting
Fracture communication between core holes while
grouting was observed. An increase of water flow
from Ml was evident while grouting M2 (M2 was
grouted prior to Ml). Similarly, blue water was
observed coming from the Wl weep hole in the
Wl drift while grouting Ml. A flow of grout and
water from previously grouted core holes M3, M4,
and M5 was observed while grouting M9. The
Wl rib leaked grout while grouting the interval of
M8 from the face to 42 ft (Table 4-8). Finally,
grout was observed to be leaking from the Wl
weep hole after grouting MJ10.
24
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Miller Mine Hole M-1
Average Lugeon and Rock Fracture Values For Diorite
5.8 Fractures/Foot
Lugeon
i/alue of 8
4.3 Fractures/Foot
Lugeon Value of 77
2.4
Fract/Foot
19 Fractures
85 Fractur
;s
173
ractures
DTotal Fractures
DAverage Lugeon Value
Grout Take (Gallons)
FacetoTD = (15)@52psi
Lugeon Value of 176
80 100 120 140
Average Lugeon and Fracture Values
Figure 4-1. Plot of the average Lugeon and fracture density for grout hole Ml.
Miller Mine Hole M-2
Average Lugeon and Rock Fracture Values For Diorite
4.3 Fractures/Foot
Lugeon Value of 12
Total Fractures; Avg.
Fractures/Foot
D Average Lugeon Value
5.5 Fractures/Foot
Lugeon Value of 61
Grout Take (Gallons)
Face to 32 Feet = (10.2) @ 60 psi
32FeettoTD=(19.9)@85psi
4.2 Fractures/Foot
Lugeon Value of 66
Average Lugeon and Fracture Values
Figure 4-2. Plot of the average Lugeon and fracture density for grout hole M2.
25
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Miller Mine Drill Hole M-3
Average Lugeon and Rock Fracture Values For Diorite/Silicious Clays tone
6.8 Fractures/Foot 68 Fractu
Lugeon Value of 51
4.8 Fractures/Foot 1 72 Fr
Lugeon Value of 46
3S
ictures
Total Fractures; Avg. Fractures/Foot
D Average Lugeon Value
Grout Take (Gallons)
Face to 53
53 Feet to T
"eet = NA
148 Fractures
4.8 Fractures/Foot
Lugeon Value of 55
4.7 Fractures/Foot
56 Fractures
Lugeon Value of 112
- 53-68
0)
Average Lugeon and Fracture Values
Figure 4-3. Plot of the average Lugeon and fracture density for grout hole M3.
Table 4-1. Rock Core and Drill Data
Core Hole _ ..
,-. , Drill/
(ft above -. , ,1
v.lr. Method
sill)
Ml
M2
M3
M4
M5
M6
M7
M8
M9
MJ10
Hagby
Hagby
Hagby
Hagby
Hagby
Hagby
Hagby
Hagby
Hagby
Jackleg
Core Size2
0-10
HQ
HQ
HQ
BQ
BQ
BQ
BQ
BQ
BQ
NA
10-TD
NQ3
NQ3
NQ3
AW34
AW34
AW34
AW34
AW34
AW34
NA
TD3
(ft)
70
70.5
78
77.2
76.3
72
57
58
58
7
Core Recovery
Total
Core (ft)
66
66
76
77.2
76
71
57
57
57
NA
%
94
94
97
100
99
99
100
99
99
NA
Mean
RQD4
28
34
27
37
34
37
36
30
18
NA
Bearing
N53E
N70E
S85E
S81E
S89E
N85E
N90E
N70E
N67E
S88E
Inclination _
_ , Core
Degrees from _ . , , ,6
TT . ,s Orientated
Horizontal
+13
+8
+10
+9
+10
+10
-8.5
+1
+6.5
+2
Yes
Yes
Yes
No
No
No
No
No
No
NA
1 Hagby 1000 core drill; Gardner-Denver, Model 83 Jackleg pneumatic-percussive drill
2 HQ core diameter 2.500 inches, hole diameter 4.625 inches; BQ core diameter 1.432 inches, hole diameter 2.965 inches; NQ3
core diameter 1.875 inches, hole diameter 2.980 inches; AW34 core diameter 1.062 inches, hole diameter 1.890 inches
3 Cored TD (total depth) except for MJ10
4 RQD (rock quality designation) is a modified core recovery percentage in which all pieces of sound core over 4 inches in length
are summed and divided by the length of the core run (Appendix C)
5 Inclination and bearing measured with Brunton Compass
6 Orientated core was recovered using the clay-impression method (Appendix C)
26
-------
Table 4-2. Comparison of Downhole Camera Survey Data
Hole Number Brunton
Ml
M2
M3
Bearing
N53E
N70E
S85E
Survey at Face
Inclination
(Degrees)
+13
+8
+10
Camera Survey
Bearing
N58E
N71E
S85E
Inclination
(Degrees)
+13
+7
+10
Depth-Ft
(TD)
70
70.5
78
Table 4-3. Survey - Drill Point Location
Drill Point
Ml
M2
M3
M4
M5
M6
M7
M8
M9
Northing
11547.8
11547.0
11544.6
11543.8
11545.2
11545.8
11545.3
11530.6
11531.5
Easting
11244.9
11245.7
11247.7
11248.1
11247.4
11247.1
11247.9
11257.4
11256.9
Elevation
6306.4
6305.8
6306.0
6305.7
6306.1
6306.1
6304.5
6307.2
6307.9
Ft Above Sill
5.3
4.7
4.9
4.6
5.0
5.0
3.4
6.1
6.8
Table 4-4. Water-Bearing Fractures (Intervals)/Communication
Core Hole No.
Water-Bearing Intervals
GPM
Core Hole Communication
During Drilling
Ml
M2
M3
M4
M5
M6
M7
M8
M9
M4
NA
43.8-48.7
53-58
58-63
68-73
23-31
45.5-46.5
NA
43-44
52-57
48-53
43-58
31.4-32.4
23-31
46
NA
0.1
2.4 to 4.1 total
0.2
1.5
4
1
NA
Approximately 2
1.5
2-3
1.1
3-4 gpm
Ml and M2 (M2 drill water
from Ml)
Water from Ml while
drilling M2 at 29 ft
Drilling fluid in the Wl drift
Drilling fluid in the Wl drift
Water flowing from back
after drill skimmed and
broke into the Wl drift
At 55.9-56.1, observed grout
from M3; blue dye seen at
37.8 in M5, likely from M4
At 37. land 43.5-43.8,
observed grout likely from
M4
At 42.4, observed grout from
M8/M6
At 57.6, observed grout from
M3; water flow increased
through back (M4 core hole
skimmed Wl back) while
drilling M5
27
-------
Table 4-5. Average Fracture Hydraulic Conductivity Values (cm/s) from Water Tests (K); Based on Lugeon Values
Core Hole Ml
Interval
(Ft)
10-18
18-38
40-70
K
2.29xlO'3
l.OOxlO'3
1.04X10'4
Core Hole M2
Interval
(Ft)
12-32
32-52
52-70.5
K
8.58X10'4
7.96xlO'4
l.SOxlO'4
Core Hole M3
Interva
KFt)
10-22
22-53
53-68
68-78
K
1.45X10'3
7.09xlO'4
5.98xlO'4
6.63xlO'4
Core Hole M4 Core Hole M8
Interval
(Ft)
32-77.2
K Interva K
KFt)
1.69xlO'4 42-58 1.82X10'4
Table 4-6. Grout Properties
Grout
Uncured
Cured
Viscosity
500-700 cps at
77 °F
NA
Color
Opaque-amber
Pale brown
Density
8.75-9.17 Ibs/gal
8.75-9. 17 Ibs/gal
Toxicity
Non Toxic
Non Toxic
Tensile Strength
NA
89psi
Table 4-7. Grout Injection Data
Core Hole Grout Injection Amount of Material
No. Intervals (Ft) Injected (gal)
Ml
M2
M3
M4
M5
M6
M7
M8
M9
Total
Face - 70
32 - 70.5
Face - 32
53-78
'Face - 53
32-77.2
32-76.3
Face - 72
Face - 57
42-58
Face - 42
Face - 58
Grout
15
19.9
10.2
24.7
NA
18
15
20
18
13.5
2
17.9
174.2
Water
9.5
5
7
5
NA
15
15
23
20
14.8
1.7
16.6
131.6
Average Time Required Date of
Injection For Injection Injection
Pressure (psi) (minutes)
52
85
60
160
NA
100
80
60
60
80
NA
80
8
7
6
12
NA
18
14
12
10
11
1
7
10/5/99
10/4/99
10/5/99
10/5/99
10/8/99
10/8/99
10/13/99
10/13/99
10/18/99
10/20/99
10/20/99
M3 interval face - 53 not grouted; grout came around packer from interval 53-78 and came to face
28
-------
Table 4-8. Water Test/Grout Communication Between Core Holes
Core Hole No.
M3
M2
Ml
M4
M5
M6
M7
M8
Water Injection
Intervals/Stages (ft)
10-22
22-53
53-68
68-78 blue dye
53-78 blue dye
30-78
12-32
32-52
52-70.5
10-18
18-38
42-70
32-77.2
42-58
Flow of Injected Water Grout Injection
Observed From Intervals/Stages
M2
M2
M3 flow ceased once 53-face
packer set at 53
53-78
Ml andM2
Face-32
32-70.5
Face-70
M2
Dye not observed in Ml ,
M2, and M3; dye
observed in the Wl drift;
flow from Wl back
increased
32-76.3
Face-72
Face-57
Wl weep hole flow 42-58
Grout/Water
Observed From
Water from Ml
Blue water - Wl
weep hole
32-77
Grout leaking from
back & rib of the Wl
drift; used polyfill
Increase in the Wl
M9
MJ10
increased
Face-42
Face-58
0-7
weep hole flow
Wl rib leaked grout;
plugged leaks with
polyfill
Grout/water from
M5, M4, M3
Plugged Wl weep
hole
29
-------
5. Phase II - Field Emplacement
As a result of the 1999 grout injection, a majority
of the flow into the Wl drift at the Miller Mine
was eliminated. However, the grout injection had
only eliminated flow that entered the mine from
either above or below the mine workings. Due to
the small size of the Wl drift (4 ft by 6 ft), it was
impossible to place the Hagby core drill in the
drift. Flow was visibly entering the mine from the
northeastern rib of the underground mine workings
as drips and weeps. A decision was made that
additional grout injection work would be
performed in an effort to eliminate flow into the
Wl drift. The work was scheduled for April 2001,
and a Jackleg drill was used to access the Wl drift.
5.1 Preliminary Design
A Gardner-Denver Model 83 Jackleg drill was
used to drill the grout holes in the Wl drift (Figure
5-1). Geological information gathered during the
October 1999 grout injection was used to map and
design the second field implementation of grout.
Five areas within the Wl drift were designated for
grout injection; the areas were denoted as Zone A,
B, C, D, and L13 and were located 18, 12, 6.9, 5
and 0 inches from survey point LI3, respectively.
The grout holes were drilled upward in a radial
pattern at a shallow angle to the back of the
workings (between 10 and 16 degrees) at each
zone location. Holes were drilled and grouted on
an alternating pattern to reduce communication
between grout holes.
Initially, all grout holes were drilled to a depth of
8 ft and had a 1.75-inch ID. Some holes were
extended in the field, but this decision depended
on the fracture density, flow, and communication
between the core hole and the Wl workings. Once
a grout hole was drilled, the flow from each hole
was denoted; mechanical packers were then placed
in each hole to reduce the amount of water
entering the mine workings. Zone A was the first
scheduled location drilled and grouted, followed
by Zone B (Appendix A, Plate 5).
The same grout materials (HA Combi grout and
the HA Flex Cat Accelerator) were used for the
April 2001 grout injection phase (Ref 1). The
amount of catalyst was varied depending on the
number of holes to be grouted per grout run, water
temperature, and amount of water flowing from
each grout hole. The amount of grout varied for
each grout run. Consideration was taken into
account for the grout hoses, grout wasted between
setups, and length of the hole (1 gallon grout per
an 8-ft grout hole).
The activated grout containers and the grout
injection equipment were transported to the grout
stations using a small Kubota tractor (Figure 5-2).
Once the grout equipment was set up, the grout
was poured into a 30-gallon grout tank and
activated using the catalyst. Grout injection began
using a Moyno screw pump powered by a
generator. Grouting pressures ranged from 60 to
300 psi, depending on the tightness of the fracture.
5.1.1 Geology and Drift Segments (A, B, C,
D, and LI 3)
Due to the limiting size of the Wl drift workings
(4 ft by 6 ft), a Jackleg drill was used to drill the
grout holes. The cuttings were not collected since
the workings were so confining, allowing only one
individual to work in the Wl drift. As a result, the
geology of each hole was not defined.
The geology mapped during the 1999 grout
emplacement showed that the majority of the
water entering the mine was either associated with
the contact zones between the Tqd and the
silicified limey shales of the Precambrian Newland
Formation or with the fracture/fault zones
associated with the MMRF. In the Wl drift, there
were select areas that produced water. These areas
were mapped before and after the first grout
emplacement, and five distinct zones were selected
for further grout emplacement using Jackleg
drilled short holes. The five zones were as
follows.
30
-------
Zone A - At zone A, water influx into the
underground mine workings was at the contact
between the quartz diorite and the silicified limey
shales. At this zone, water drips from the roof and
the south mine wall. Fractures were at a lower
angle (61 to 53 degrees from vertical). Lower
angle fractures were not very tight (compressed)
and usually produced or allowed increased water
movement compared to steeper angle fractures and
faults.
Zone B - Zone B was located at the contact
between the quartz diorite and the silicified limey
shale; however, Zone B was much tighter and the
fractures are near vertical to the mine workings.
This zone had minimal influx, and the influx was
from drips in the wall that dampened the wall
surface.
Zones C, D, and L13 - Zones C, D, and L13 were
all associated and located in the MMRF zone
(Appendix A, Plate 5). In the area of survey point
LI 3, the mine workings followed the strike of the
fault and the zone of numerous (less than 1 inch)
quartz stringers; however, the quartz stringers
were discontinuous and low grade, which led to
the abandonment of the Wl drift. The rock in this
area was very broken, and the northeastern and
eastern rib of the Wl drift was inaccessible to the
1999 grout emplacement due to the size of the
drilling equipment used during that emplacement.
However, this area was grouted during the 2001
grout emplacement.
5.2 Summary of April 2001 Grout
Emplacement
Jackleg grout holes (1.75-inch ID) were drilled in
a radial pattern to intercept either the petrologic
contact or the fracture system associated with the
MMRF (Appendix A, Plate 5). Drilling and
grouting was initiated at Zone A and progressed to
the zones further in the Wl working.
The planned drill pattern was produced by drilling
every other hole first and then completing the
planned pattern by infill drilling between the
initial holes. The flow of water from each hole
was measured directly after each hole was drilled.
Mechanical packers were then placed in each hole
to reduce the amount of water entering the
working area (Figure 5.1). The average volume of
grout, estimated on a per foot of hole basis, was
0.23 gallons of grout per foot of hole grouted. The
HA Combi grout had an assumed expansion factor
of approximately 5 to 7 times its original
nonactivated volume. However, the expansion
factor of the grout is inversely proportional to the
confining pressure (e.g., bearing pressure in
fractures) in the system where it is being injected.
During injection, water was added so that the
grout would activate. This procedure was used
more in holes that did not produce a significant
volume of water after they were drilled than in
holes that communicated with the Wl drift
(Table 5-1).
Table 5-1 provides the date holes were drilled and
grouted, hole designation, total depth and bearing,
grout volumes, injection pressures, water volumes,
and comments. For the design of the 2001 field
injection, there were several water-bearing zones
targeted for grout injection. The zones were
designated Zones A-D and Zone L13, located
close to the survey point L13. The zones were
drilled and then grouted (Figure 5-2), and the
characteristics of each zone are as follows.
Zone A - Approximately 25 gallons of water-
activated, polyurethane grout was injected into
Zone A at pressures ranging between 200 to
330 psi. Grout holes on the west side of the
workings were tight and fairly dry (slow drips)
(Appendix A, Plate 5). However, the east side of
the workings was very fractured and soft, allowing
an increased influx of water into the mine
workings. Reducing the inflow of water to this
zone was difficult. Grout holes were drilled
having an eastern orientation, as well as a western
direction. Results show that water moved from the
grouted area to nongrouted areas.
Zone B - This zone was tight, and the fractures
were high angle. All grout holes drilled at Zone B
were dry. Only three holes were drilled and
31
-------
grouted in this area. The total grout take for the
three holes was 2 gallons.
Zone C - All grout holes drilled in Zone C had
water flowing from them. C5ALT and C7 did
intercept a high flow zone in the fracture/fault
system. These holes were placed at low angles to
the walls of the workings and drilled to intercept
the MMRF.
Zone D - Zone D was drilled from underground
survey point L13. D4, located in the east side (rib)
of the mine workings, intercepted a high flow zone
and the MMRF. Flow was greater than
3 gpm. The maximum grout injection pressures
for D4 was approximately 50 psi. The grout holes
drilled in the roof and on the west side of the
workings were dry. All holes were grouted for
closure.
Zone L13 - At underground survey point LI3,
water dripped from the roof and east side of the
mine workings. Drill holes L13A and JK10R,
drilled above the collar of hole number MJ10,
provided the most inflow (1.5 gpm) into the
workings of this group of holes The fault/fracture
system associated with the MMRF was contacted
by Jackleg drill holes at locations Zones D and
L13. Using the underground mine map (Figure
2-1) and the geological maps created from the core
drilling that occurred during the 1999 grout
emplacement, it was apparent that these holes
would contact the main water-bearing zone that
contributed to the generation of AMD in drift Wl
(Figure 5-3).
5.2.1 Grout Emplacement
Upon finalization of the April 2001 grout
emplacement, approximately 400 ft of hole had
been drilled and grouted. The total volume of FŁA
Combi grout injected was 69 gallons, and of that,
approximately 24 gallons were wasted. Wasted
grout included the gallons of grout used to fill the
grout injection lines and to recirculate the grout
through the grout tank system. Some grout was
wasted/lost when switching connections between
the mechanical packers.
On average, approximately 0.17 gallon of grout
was injected per foot of hole. However, grout
holes with increased permeability took more grout
on a per foot basis. The locations that took
increased amounts of grout were located on the
east and northeast side of the workings.
32
-------
Figure 5-1. Stabilizing the underground mine working
of the Miller Mine with roof bolts and screen using a
Gardner-Denver Model 83 Jackleg drill.
Figure 5-2. Kabota tractor that transported the grout and grout
equipment into the abandoned underground workings of the Miller Mine.
33
-------
Figure 5-3. Iron oxide stalactites in the Wl drift, which
were areas where flow through the fractures required
grouting.
34
-------
Table 5-1. Summary of the Grout Injection Sequence at the Miller Mine, W-l Drift
Date
4/16/01
4/16/01
4/16/01
4/18/01
4/19/01
4/16/01
4/16/01
4/20/01
4/20/01
4/20/01
4/24/01
4/20/01
4/20/01
4/24/01
4/25/01
4/25/01
4/17/01
4/19/01
4/17/01
4/17/01
4/17/01
4/20/01
4/17/01
4/17/01
4/20/01
4/17/01
4/19/01
4/19/01
Grout Hole
Al
A2
A3
A4
A4ALT
A5
A6
All(ABlL)
A12 (AB2V)
A13(AB3)
A14
(ARRV)
A15
(A3ALT)
A16
(A2ALT)
AC (ARRC)
ARRM
ARRM2
B2
B3
B4
Cl
C2
C3
C4ALT
C4
C5
C5ALT
C6
C7
C7ALT
C7ALT RR
(C7B)
Total
Depth
(ft)
10.5
12
8
10
10
8
6
4
5
5
6
10
10
6
6
4
17
8
15
8
17
19
15
15
17
19
19
21
Compass
Bearing
Drilled
N54E
N54E
N67E
N69E
N72E
N80E
N80E
S60W
S62W
S50W
S70W
S70W
S80W
S70W
S27W
S27W
N30E
N45E
N80E
N30E
N30E
N42E
N59E
N65E
N65E
N70E
N75E
N75E
N78E
Grout
Injection
Volume
per Hole
(gallons)
1
2
3
1.2
1
4
1
1
2.5
1.5
1.75
1
2.5
0.25
0.75
0.50
2
1
2
2
2
2
2
2.5
2.5
3
Maximum
Pressure
during
Grouting
(psi)
300
280
200
200
200
200
300
200
200
330
220
260
300
300
140
Out hole
300
150
280
200
200
Volume Water
Flowing from
the Open Hole
Drip
Dry
Dry
1 gpm
Drip
Steady drip
Small drip
Steady flow
Drip
Steady drip
Drip
Small steady
flow
Drip
No water
Drip
Dry
Dry
Dry
Dry
Dry
Drips
Small drip
Drips
Steady flow
Drip
High flow
Comments
Hit fracture at 5 ft from collar
Shallow angle drilled
Did not hit fractures
Steady flow, fracture at 6-7 ft
from collar
Located between A4 and A5
Hit fracture at 5 ft from collar
Hit fracture at 6 ft from collar
Water pushed packer out of
wall - rough ground
Drilled into right rib of drift
Drilled into right rib of drift
Drilled toward mine entrance
Located between A3 and A4
Located between A2 and A3
Drilled toward mine entrance
Drilled toward mine entrance
Drilled toward mine entrance
Dry, grouted to fill hole
Closed mechanical packer
Closed mechanical packer
No fractures hit during
drilling
Not drilled, dry area
Packer has bad valve, in roof
Drilled 3 ft over weep
Block from roof dropped, 2 ft
of hole damaged
Drilled 3.5 ft back from C5
Packer fell out of hole
Low angle hole, Miller Mine
Fault
Hit fracture at 15- 19 ft;
hard drilling past 1 9 ft
35
-------
Date
Grout Hole
4/20/01
4/17/01
4/24/01
4/17/01
4/25/01
4/25/01
4/24/01
4/24/01
Total Compass
Depth Bearing
(ft) Drilled
Grout Maximum Volume Water
Injection Pressure Flowing from
Volume during the Open Hole
per Hole Grouting
(gallons) (psi)
Comments
4/19/01
4/20/01
4/20/01
4/20/01
Dl (DMB)
D2
D3
D4
12
10
12
12
N11W
N2E
N3E
N5E
2
2
2
4
300
300
100-60
50
Dry
Dry
Dry
3gpm
Roof hole, left rib
Roof hole, center
Hit high flow zone, fault
D5
1L13H(U)
4/17/01 1L13RR
2L13H
2L13RR
2L13R
L13A
L13B
JK10R
CLR
12
15
15
8
15
15
6
3
N5E
N85E
N11E
N65E
N30E
N11E
N2E
S50W
1.5
1
Not
grouted
2
1
1.5
0.50
2
0.50
300
300
Not
grouted
200
300
250
300
220
300
zone
Drip Plugged drill, grout at 6 ft
A few drops Above MJ10, fracture at 3 ft
from collar
Dry Roof,+10 degree hole
Drops Fracture at 5 ft from collar
Drops Fracture at 5 ft from collar
Not drilled
Steady flow Lost bit and striker
Steady flow Lost half gallon at packer
Water Drilled into fault zone
Some water Drilled over weep
36
-------
6. Long-Term Monitoring and Mine Maintenance Results
6.1 Mine Maintenance
The Miller Mine project started in 1998 and was
extended through 2003. During this period, the
mine was opened with a 3-year project access
requirement. As the project progressed, there were
two technology applications rather than the
originally planned single technology application,
and the long-term monitoring was extended
through 2003. As a result, the mine was
restabilized several times; originally it was
reopened and stabilized with the thought that the
mine only needed to remain open for 3 years.
During each technology application, the mine
required stabilization, and in April 2001, there was
significant stabilization of the underground mine
workings using screens and roof bolts (Figure 5-
2). However, during the July 2001 sampling
event, it was discovered that some of the timbers
in the lower portal of the Miller Mine had
collapsed and an area inside the mine had
experienced some movement either due to
freeze/thaw events, increased water from spring
runoff and storm events, or localized seismic
movement. As a result of this finding, access to
the Miller Mine underground workings was
determined to be unsafe, and further sampling was
performed only at the Miller Mine portal (W4) and
not in the underground workings at Wl, W2, and
W3 sample ports (Figure 2-1 and Appendix A
Plate 1).
6.2 Monitoring History and Methods
For this demonstration, flow from the Wl drift
was monitored at the Wl sampling port by a small,
60-degree trapezoidal flume located at the front of
the Wl drift (Figure 1-5 and Figure 2-1). Flow
from the winze and W3 drift was monitored at the
W2 and W3 sample locations using
22.5-degree V-notch weirs. The total flow from
the Miller Mine was monitored at the portal
initially using a 22.5-degree, V-notch weir. In
addition, a volumetric bucket test was performed
at the 10-inch effluent discharge pipe. Because of
mine instability, measurements at Wl, W2, and
W3 were discontinued. The last sample taken
inside the mine was on November 8, 2001.
Monitoring was continued at the Miller Mine
portal, sample port W4, until November 2003.
The flow and water quality results are presented in
Appendix B.
6.3 Long-Term Monitoring Results
The primary criteria for the success of the project
was to reduce the cumulative volumetric flow at
the Wl sample location by 95% using flow data
from 10 months before and after grout was
injected into the Miller Mine fracture system
(Ref 2).
The first phase of grouting occurred September 24,
1999, to October 23, 1999, and the second phase
of grouting occurred April 4, 2001, to April 16,
2001. Figures 6-1 and 6-2 show the flows with the
approximate dates of grout completion.
The average flow at Wl prior to the first grout
emplacement was 5.88 gpm, and the flow after
both grout emplacements averaged 1.24 gpm, an
approximately 80% reduction.
For statistical evaluation of the flow from the Wl
drift, plotting the flow measurements [gallons per
day (gpd)] versus time (days) and the area under
the curve prior to grouting and after grouting were
compared and used to calculate the percent
reduction in flow (gallons) (Appendix F). During
the comparison, corresponding seasonal time
periods were evaluated to minimize the effect of
seasonal fluctuations in the flow comparison.
Table 6-2 provides the flow measurements taken at
the Wl drift over the duration of the project. The
flow was categorized into flow prior to, after
Phase I, and after Phase II grout injection. Note
the number of days elapsed for each time period is
listed in Table 6-2.
Data from Table 6-2 were entered into MATLAB
software as separate data strings; the area under
each curve was then calculated to determine the
37
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total volume of water that flowed through the Wl
flume during each time period (Appendix F).
Volumes between the pregrouting phase and Phase
I grout injection were compared to determine the
percent reduction in flow volume due to the Phase
I grouting. Flow values between Phase I and
Phase II were also compared to determine if the
additional grouting reduced flow. Table 6-3
provides the results from the MATLAB
calculations. Because the timeframes from the
grout phases did not match exactly, date values
after the Phase I grouting were matched with the
compared grout phase, and flow values were then
interpolated.
Table 6-3 reveals that the Phase I grouting reduced
flow volumes during the December 1999 to
September 2000 time period by 77%. The
additional grouting of Phase II did not reduce
flows any further. The flow volume during the
July to November 2001 time period increased by
21% when compared to the July 2000 to
November 2000 time period. However, flow
calculations did not consider the effect of
precipitation. During the specified time period,
the total precipitation between 2000 and 2001
increased by 34% (Ref 6). This increase in
precipitation was not taken into account in the
flow calculations at the Miller Mine.
Considering that the trapezoidal flume at Wl had
an accuracy of+/- 5%, the flow reduction ranges
were calculated by using the +5% maximum and
-5% minimum flow values for each phase. Table
6-4 shows that the flow from the pregrout phase to
after Phase I grouting was reduced by 75.1% to
79.6%. The flow difference between the Phase I
grouting and the Phase II grouting was increased
by 9.8% to 34.1%.
Long-term flow and monitoring results indicate
that the grouting program has reduced the flow
from the Miller Mine adit (W4), as well as the
concentrations of dissolved metals. Flow and
analyses of dissolved metals were recorded for
4 years after completion of the primary grouting
program in 1999.
The influent flow (cumulative monthly flows at
Wl and W2) and the effluent flow at the Miller
Mine adit (W4) were compared and evaluated
(Figure 6-1). Results indicate that the percent
reduction in flow at the Miller Mine adit (W4)
ranged between approximately 50% and 84%
depending on the season (Figure 6-1).
Prior to the application of the source control
technology, flow results and visual observation
revealed that most of the flow into the workings
occurred in the Wl drift (Figure 6-2). After
grouting, flow in Wl was reduced; however, flow
recorded from the W2 and W3 drifts indicate that
those workings were not affected by the grout
emplacement even with the close proximity of the
workings.
The last flow measurements taken inside the
Miller Mine workings were on November 8, 2001.
On that date, 53% of the flow was from the Wl
drift and 47% was from the W2 drift. As the flow
at W4 increased from 1.9 gpm to approximately
3.0 gpm in September 2003, it is unknown where
the increase in flow originated because access to
the underground mine workings was restricted.
However, since there was an increase in
precipitation of approximately 34% from the prior
year, then part of the increased flow could be a
result of increased precipitation. The effect of the
increased precipitation on the in-mine flow regime
is not known.
6.4 Water Quality Results at the Miller
Mine
The secondary criteria for success at the Miller
Mine included improving the quality of the water
discharging the adit; decreasing the dissolved
metals concentrations, and increasing the pH of
the water discharging from the Wl drift and the
mine portal at sample port W4. The analyzed
dissolved metals were Al, arsenic (As), cadmium
(Cd), Cu, Fe, Pb, Mn, Ni, Zn, and silver (Ag);
sulfate concentrations were also analyzed.
Because of the wide range in concentration values,
the dissolved metal concentrations and metals
loading graphs were plotted using a logarithmic
38
-------
scale (Figure 6-3). The dissolved metal
concentrations for As and Ag were at or below the
instrument detection limit (IDL) and will not be
discussed in detail.
6.4.1 Water Quality and Metals Loading at
Wl
During evaluation of the dissolved metal
concentrations from July 1998 to October 2001, it
was determined that application of the technology
did not significantly affect the water quality in the
Wl drift. It was thought that by grouting the
exposed fracture surfaces (much of which is
sulfide material) and eliminating water or oxygen
(required for the formation of AMD) that the
dissolved metals concentrations in the Wl drift
would be reduced. Water quality results from Wl
indicate that there was a slight decrease in pH
from 6.08 to 5.95 and a minimal increase in
dissolved metals concentration for Al, Cd, Cu, Fe,
and Pb (Table 6-5). Conversely, a decrease in the
concentrations of Zn, Mn, Ni, and sulfate were
recognized (Figure 6-3 and Table 6-5).
However, due to the reduction in flow in the Wl
drift, the metal-loading rates (both dissolved and
total) were reduced significantly, even though the
water quality (as concentration) did not fluctuate
(Figures 6-4 and 6-5). Average dissolved metal
load reductions of greater than 80% were
calculated for Cd, Al, Zn, and Fe. Reductions of
greater than 50% were recognized for the Mn, Pb,
Ni, and Cu dissolved loads (Figure 6-4). Specific
metals-loading reductions calculated for the Wl
drift (Figure 6-5) indicate:
- Fe load was reduced from 7.50 pounds per
day(lb/d)to0.121b/d;
- Mn from 0.45 Ibs/d to 0.09 Ib/d; and
- Zn from 0.08 Ibs/d to 0.004 Ib/d.
It should be noted that the data used in the
calculation of load reduction were taken from the
same season (fall-early winter) as the flow of
water and the concentration of metals in that water
(both load calculation variables) are known to be
strongly effected by seasonal variations.
6.4.2 W4 Miller Mine Portal Water Quality
and Metals Loading
The water quality of the water discharging at the
Miller Mine portal (W4) improved as a result of
the technology applications. The water quality
improvement is directly related to the reduction of
influx into the Wl drift. The metal concentrations
from Wl are being diluted as the water from Wl
merges with water from the W2 drift. The amount
of dissolved metal in the portal discharge
significantly dropped after each application of the
technology (Figure 6-6). Iron and Zn
concentrations were reduced the most when
comparing 1999 pregrout and 2003 postgrout
analytical results. The dissolved concentrations
for Pb did increase; however, due to the decrease
in flow, Pb loads were reduced by 50% on average
(Figure 6-7).
As calculated, the average percent reduction of
dissolved metal loads was greater than 50% for all
metals analyzed. Both Zn and Fe loading was
reduced as much as 90%, and the Phase II grouting
assisted with further reduction in the metals
loading (Figure 6-8). The most significant
reduction in metals loading at the Miller Mine
portal, located 600 ft from the Wl drift, include:
- Fe from 3.05 Ib/d to 0.02 Ib/d;
- Mn from 0.46 Ib/d to 0.09 Ib/d; and
- Zn from 0.06 Ib/d to 0.004 Ib/d.
Most of the analyzed dissolved metals load was
reduced by an order of magnitude or more.
39
-------
Inflow and Outflow at the Miller Mine
Phase I Grout
Injection 10/99
W-1, W-2, and
W-3 added
together
W-4 Manual
Flow
Measurement
(gpm)
NOTE: W-4 flow
is zero during the
winter months
Date
Figure 6-1. Inflow (Wl, W2, W3 added together) and outflow (W4) measurements compared over time to
evaluate losses and fluctuations due to grout injection.
Miller Mine Flow Measurements
Note: Grout Injection occurred between:
Phase I - 9/24/99 to 10/23/99
Phase II-4/1/01 to 4/16/01
W-4 = 0.0 during winter months and is the
portal measurement
DW-3 Manual Flow Measurement (gpm)
W-4 Manual Flow Measurement (gpm)
DW-2 Manual Flow Measurement (gpm)
W-1 Manual Flow Measurement (gpm)
Figure 6-2. Miller Mine flow measurement taken at the respective locations throughout the mining system.
Note that the Miller Mine became unstable, and measurements for Wl, W2, and W3 were discontinued in
November 2001.
40
-------
Miller Mine Water Quality at W1 Drift
1000000
Note: tjout injection occurrea Detween
Phase I -9/24/99 to 10/23/99
Phase II-4/1/01to4/16/D1
Date
Figure 6-3. Miller Mine water quality results for the Wl drift, the area where most of the dissolved metals
were originating. The water quality results reflect the dissolved metals concentrations at Wl micrograms per
liter.
Average Percent Reduction at the Miller Mine, W1 Drift
-------
Miller Mine Loading Rates at W1 Drift
Note: Grout Injection occurred between:
Phase I - 9/24/99 to 10/23/99
Phase II-4/1/01 to 4/16/01
W4 = 0.0 during winter months
Figure 6-5. Metals loading at the Wl drift sample port located in the underground workings of the Miller
Mine.
Miller Mine Water Quality at the Portal
(W4)
100000
Note: Grout Injection occurred between:
Phase I-9/2*99 to 10/23/99
Phase II-4/1/01 to 4/16/01
Date
Figure 6-6. Water quality at the Miller Mine portal (W4 sample port).
42
-------
100% -
90% -
80% -
70% -
c
o
= 60% -
3
1 50% -
"c
0 40% -
0>
0.
30% -
20% -
10% -
0% -
Average Percent Reduction at the Miller Mine Portal
(W4)
~
~
-
, .
-
-
As Cd Cu Ni Pb Al Zn Win Fe
Metals Analyzed
DAfter Phase I
After Phase II
Grout -2001
[ULast Sampling
Event - 2003
Figure 6-7. The average percent reduction in metals loading at the Miller Mine portal after grout injection.
Miller Mine Loading Rates at the Portal
(VW)
Note: Grout Injection occurred between
Phase I - 9/24/99 to 10/23/99
Phase II-4/1/01 to 4/16/01
Figure 6-8. Metals loading for the Miller Mine portal (sampling results from port, W4).
43
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Table 6-1. Wl Drift Recorded Flow Data Taken for the Duration of the Technology Demonstration
Date
12/10/98
1/8/99
2/8/99
3/3/99
5/3/99
7/8/99
8/5/99
9/23/99
11/10/99
12/14/99
2/25/00
GPD
7,488
7,488
7,488
5,184
5,184
8,856
10,354
15,696
2,160
2,160
2,007
Date
4/20/00
6/1/00
9/15/00
9/27/00
11/30/00
12/28/00
2/1/01
7/23/01
9/25/01
11/8/01
GPD
2,002
1,536
1,482
1,499
1,482
1,512
2,160
1,051
2,002
1,469
Table 6-2. The volume of Flow from the Miller Mine, Drift Wl, Prior to and After
Each Phase of Grouting on a Gallons per Day Basis
Prior to Grouting
Date
12/10/98
1/8/99
2/8/99
3/3/99
5/3/99
7/8/99
8/5/99
9/23/99
Day
0
29
60
83
144
210
238
287
GPD
7,488
7,488
7,488
5,184
5,184
8,856
10,354
15,696
After Phase I Grouting
Date
11/10/99
12/14/99
2/25/00
4/20/00
6/1/00
9/15/00
9/27/00
11/30/00
12/28/00
2/1/01
Day
0
34
107
162
204
310
322
386
414
449
GPD
2,160
2,160
2,007
2,002
1,536
1,482
1,499
1,482
1,512
2,160
After Phase H Grouting
Date
7/23/01
9/25/01
11/8/01
Day
0
64
108
GPD
1,051
2,002
1,469
44
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Table 6-3. Flow Reductions at the Wl Drift
Grout Phase
Pregrouting
Post Phase I
Grouting
Post Phase I
Grouting
Post Phase II
Grouting
Wl Flow
Time Frame
December 1998 to
September 1999
December 1999 to
September 2000
July 2000 to November 2000
July 2001 to November 2001
Calculations
Total Gallons
2,282,036
514,481
143,490
174,058
Percent Reduction of Flow
77% reduction between Pre-
Grouting and Phase I
21% increase of flow between
Phases I and II
Table 6-4. Wl Drift Flow Reductions with Error Ranges
Grout Phase Time Frame Flow Volume +5% Volume -5% Volume
Pregrout 12/98 - 9/99
Phase I 12/99 - 9/00
Phase I 7/00-11/00
Phase II 7/01-11/01
2,282,036 2,396,138
514,481 540,205
143,490 150,665
174,058 182,761
2,167,934
488,757
136,316
165,355
Max % reduction Min % reduction
79.6% 75.1%
-9.8% -34.1%
Table 6-5. Water Quality Results for the Miller Mine after Both Phases of Field Emplacement of the Source Control
Technology
MWTP - Field Sampling Comparison
Site: Miller Mine
MWTP, Activity IE, Project 8
Sample Location: Comparison of All Sample Locations
Sample Date:
Sample Location:
Sample Matrix:
Mine Inflow or Outflow
pH (su)*
Sulfate*
Dissolved Metals (mg/L,
ppm)
Al
As
Cd
Cu
Fe
Pb
Mn
Ni
Zn
7/8/1999
Wl
Water
Inflow
6.08
828
0.145
0.025
0.003
0.002
81.600
0.023
5.970
0.097
0.915
11/8/2001
Wl
Water
Inflow
5.95
648
0.185
0.037
0.0048
0.003
95.200
0.051
5.490
0.069
0.500
11/8/2001
W2
Water
Inflow
6.4
941
0.028
0.037
0.0048
0.003
17.600
0.051
0.956
0.020
0.0949
7/8/1999
W4
Water
Outflow
6.58
539
0.010
0.025
0.003
0.002
18.600
0.023
3.220
0.046
0.388
11/8/2001
W4
Water
Outflow
6.65
510
0.028
0.037
0.0048
0.003
12.100
0.051
3.520
0.041
0.183
11/20/2003
W4
Water
Outflow
6.75
453
0.031
0.036
0.0068
0.0014
0.744
0.058
3.310
0.043
0.151
MCL & Secondary
MCL, mg/1
6.5-8.5
250
0.050 - 0.200*
0.050 until 1/06
0.005
1.000*
0.300*
0.015
0.050*
0.088**
5.000*
* Secondary MCL
** Aquatic Life Standard for Freshwater
45
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7. Summary of Quality Assurance Activities
7.1 Background
The following is a summary of the quality
assurance (QA) activities associated with MWTP
Activity III, Project 8, Underground Mine Source
Control. Analytical samples and field data were
collected according to the schedule outlined in the
approved project-specific QAPP. All field and
laboratory data available was evaluated to
determine the usability of the data. Phase I critical
analyses were surface water flow rate manually,
surface water flow rate by weir, field pH, and
dissolved metals (Al, As, Cd, Cu, Fe, Pb,
magnesium (Mg), Mn, Ni, K, Na, and Zn). Phase
III critical analyses were surface water flow rate
manually and surface water flow rate by weir.
Phase IV critical analysis was surface water flow
rate manually. A critical analysis is an analysis
that must be performed in order to determine if
project objectives were achieved. Data from
noncritical analyses were also evaluated.
7.2 Project Reviews
An external technical systems audit of the project
field activities and the HKM Laboratory was
performed by Science Applications International
Corporation, subcontractor to EPA, on September
27, 2000. There were no findings, two
observations, and one additional technical
comment identified during the audit.
The observations included deviations from the
quality objective for surface water flow manually
and inadequate calibration verification checks on
the pH meter. Amendments were made to the
QAPP to correct the observations. The additional
technical comment concerned the difference (30%
or greater) between the manual head reading on
the weirs and flumes and the pressure transducer
reading. The reason for the inaccurate readings
from the pressure transducers was investigated.
The QAPP was amended to state that manual
readings would be the sole head measurements.
7.3 Data Evaluation
Data that were generated throughout the project
were validated. The purpose of data validation is
to determine the usability of data that were
generated during a project. Data validation
consists of two separate evaluations: an analytical
evaluation and a program evaluation.
7.3.1 Analytical Evaluation
An analytical evaluation of all data was performed
to determine the usability of the data that were
generated by HKM Laboratory for the project.
Laboratory data validation was performed using
USEPA Contract Laboratory Program National
Functional Guidelines for Inorganics Data Review
(Ref 7) as a guide. The data quality indicator
objectives for critical measurements were outlined
in the QAPP and were compatible with project
objectives and the methods of determination being
used. The data quality indicator objectives were
method detection limits (MDLs) accuracy,
precision, and completeness. Control limits for
each of these objectives are summarized in Tables
7-1 and 7-2. The quality control (QC) criteria
were also used to identify outlier data and to
determine the usability of the data for each
analysis.
Measurements that fall outside of the control limits
specified in the QAPP, or for other reasons were
judged to be outliers and were flagged
appropriately to indicate that the data were judged
to be estimated or unusable. All data requiring
flags are summarized in Table 7-3.
7.3.2 Program Evaluation
Program evaluation includes an examination of
data generated during the project to determine that
all field QC checks were performed and within
acceptable tolerances. Program data that were
inconsistent or incomplete and did not meet the
QC objectives outlined in the QAPP were viewed
as program outliers and were flagged appropriately
to indicate the usability of the data.
46
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7.3.2.1 Surface Water Flow Rate
Surface water flow rate was measured with weirs,
a flume, and manually. The flume/weir water
depth was measured with pressure transducers and
staff gauges installed at each weir and flume.
During the field activities, it was discovered that
the pressure transducer measurements and staff
gauge measurements differed by as much as 30%
or greater. It was apparent that the pressure
transducers did not stay constant; therefore, the
manual reading of the staff gauges became the sole
measurement for flume/weir water depth.
As outlined in the QAPP, manual checks of
surface water flow rate were performed at each
available weir location during sampling events at
the site. Manual flow rate measurements were
performed using a graduated cylinder and a
stopwatch. Water was collected in a 1-L
graduated cylinder. The time necessary to fill the
graduated cylinder was measured with a
stopwatch. The flow rate was calculated by
dividing the volume collected by the time elapsed.
The test was then repeated to ensure the precision
of the method. The results of the duplicate test
were to be within ą 200 milliliters per minute
(mL/min) for the flow rate measurement to be
valid. If the results of the duplicate test were not
within ą 200 mL/min, a third manual flow rate
check was performed and compared to the other
flow rate measurements. If the third measurement
was within ą 200 mL/min of one of the first two
measurements, an average of the two comparable
measurements was used for reporting purposes.
The surface water flow rate measurements were
obtained in accordance with the procedures
outlined in the QAPP. No surface water flow rate
data was judged to be outlier.
7.3.2.2 pH
The pH measurements were a critical analysis for
Phase I. The pH measurements were collected
using a pH meter with automatic temperature
compensation capable of measuring pH to ą 0.1
pH units. The pH probe was calibrated daily using
two fresh buffer solutions (4 and 7). Meter
calibration was verified following initial
calibration and every 10 samples using a third
buffer solution (pH 5) within the calibration range.
The QAPP required that the calibration
verification standard be within 0.2 pH units of the
true value and that duplicate pH readings be
conducted every 10 samples or 1 per batch,
whichever is more frequent. The calibration
verification standard was within acceptable limits.
The QAPP also required that a sample duplicate be
conducted every 10 samples or 1 per batch,
whichever is more frequent. Sample duplicates
were collected for each batch, but three of the
samples did not have a pH measurement recorded
in the logbook or field notes; therefore, the pH
data for those three sampling events was flagged
"J" as estimated. A summary of the flagged data is
presented in Table 7-3.
7.3.2.3 Dissolved Metals
Dissolved metals analysis was a critical analysis
for Phase I. Aqueous samples were collected from
the four sampling locations during each sampling
event, as well as a duplicate sample from a
predetermined sampling location. Sampling
procedures for the collection of the aqueous
samples outlined in the QAPP were followed. The
samples were taken to HKM Laboratory for
analysis by inductively coupled plasma emission
spectrometer (ICP-ES). No dissolved metals data
was judged to be outlier.
7.4 Summary
Although field log sheets were used to collect
specific data during sampling events, critical
activities should be documented in the field
logbooks. In addition, one logbook should be
used for recording sampling activities; this project
used nine logbooks. The information provided in
the logbook needs to be expanded and better
organized so that anyone reviewing the logbook
can clearly understand what occurred at each
sampling event. The importance of logbook
protocol should be reiterated to sampling
personnel.
47
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Table 7-1. QA Objectives for Accuracy, Precision, MDL, and Completeness
Measurement
Flume/weir water depth
Surface water flow rate
(weir/flume)
Surface water flow rate
(manual)
PH
Units
Inches
Gpm
mL/min
S.U.
MDL
0.03
1
200 mL/min
2
Precision1
N/A3
N/A3
+ 200
mL/min5
+ 0.26
Accuracy
+ 5%4
+ 5%4
N/A4
+ 0.27
Completeness2
95%
95%
95%
95%
Dissolved metals
mg/L
See Table 6-2 < 20% relative 75%-125%
percent spike recovery
difference
(RPD)
95%
'Precision will be determined by the RPD of duplicates, unless otherwise indicated.
Completeness is based on the number of valid measurements, compared to the total number of samples.
Duplicate measurements of field process measurements will not be taken. All equipment is calibrated against National
Institute of Standards and Technology (NIST) traceable standards.
4Accuracy of weirs/flumes will be ensured by installing flumes and weirs according to Standard Operating Procedure (SOP)
H6-6 and avoiding installation locations that could adversely affect weir/flume accuracy (approach conditions do not allow
uniform velocity distribution, damage to weirs or flumes, and changes in weir or flume dimensions). In addition, manual
flow rate measurements will give an indication of whether the weirs and flumes are returning reasonable flow rate
measurements.
'Precision of manual surface water flow rate measurements will be determined by the absolute difference between
consecutive measurements.
Precision of pH measurements will be based on the absolute difference of duplicate readings.
7Accuracy of pH measurements will be based on absolute difference of reading compared to standard buffer solution.
Table 7-2. Instrument Detection Limits (IDLs) for ICP Analysis of Dissolved Metals
Analyte
Al
As
Cd
Calcium
Cu
Fe
Pb
Mg
Mn
Ni
Potassium
Sodium
Zn
1 These IDLs are considered
2 Based on National Primary
IDL (ppm)1
0.017
0.026
0.004
0.010
0.002
0.014
0.024
0.009
0.003
0.015
0.017
0.007
0.009
MDL (ppm)2
0.05-2.0
0.05
0.01
N/A
1.3
0.3
0.05
N/A
0.05
N/A
N/A
N/A
5.0
sufficiently low for the characterization of this site.
and Secondary Drinking Water Regulations.
48
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Table 7-3. Summary of Flagged Data for Activity in, Project 8
Date of
Collection
Sample ID
Analysis
Quality Criteria
Flag
Comment
8/26/1998
10/16/1998
11/17/1998
MMW1
MMW2
MMW3
MMW4
MMW1
MMW2
MMW3
MMW4
MMW1
MMW2
MMW3
MMW4
PH
PH
PH
Duplicate every 10 J
samples or one per batch,
whichever is more
frequent
Duplicate every 10 J
samples or one per batch,
whichever is more
frequent
Duplicate every 10 J
samples or one per batch,
whichever is more
frequent
No duplicate pH samples
recorded in field logbook or
field notes. The associated
samples should be flagged
"J" as estimated.
No duplicate pH samples
recorded in field logbook or
field notes. The associated
samples should be flagged
"J" as estimated.
No duplicate pH samples
recorded in field logbook or
field notes. The associated
samples should be flagged
"J" as estimated.
Data Qualifier Definition:
J - The measurements are estimated.
49
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8. Conclusions and Recommendations
The MWTP, Activity III, Project 8, Underground
Mine Source Control Demonstration Project,
reduced influx by injecting a water-activated
polyurethane grout through the water-bearing
fracture systems associated with the Miller Mine
fault zone and secondary fractures associated with
the contact of Belt sedimentary rocks and intrusive
igneous rocks. During the technology
emplacement, the limit of grout communication
between core holes was controlled by the
reduction in hydraulic conductivity of the fracture
system and the viscosity of the polyurethane grout.
Grouted fracture hydraulic conductivity is a
function of not only the fractured medium but also
of the grout density, viscosity, and rate of
reactivity.
Two fracture-shear systems were identified and
mapped from core data. This included the
fracture-shear systems associated with the
movement on the MMRF and fractures oriented at
high angles to that structure. It is believed that the
MMRF-related fracture-shear system provided the
principal water path for AMD into the Wl drift.
Moreover, the two fracture systems are
interconnected in the immediate vicinity of the Wl
drift and act together to conduct AMD charged
groundwater into the Wl drift.
Grout injection into the rock fracture system and
associated reduction of AMD was observed in the
Wl drift and W4 (adit) following completion of
the 1999 Stage I and 2001 Stage II grouting
programs. The Stage I application was responsible
for reducing the majority of the discharge from
W4 and the Wl drift. The primary criteria for the
success of the demonstration project was to reduce
the cumulative volumetric flow at Wl by 95%
using flow data from
10 months prior to and 10 months after both grout
injections into the Miller Mine fracture system.
However, average calculated flow at the Wl
sample location, prior to the first grout
emplacement in 1999, was 5.88 gpm, and the flow
after both grout emplacements averaged 1.24 gpm,
providing an approximately 77 +/- 5% reduction.
Approximately 75% of the flow reduction at the
Miller Mine was achieved within 1 month after the
completion of the primary grouting program in
1999. Long-term flow and monitoring results
indicate that the grouting program reduced the
flow from the Miller Mine adit (W4), as well as
the concentrations of dissolved metals. Flow data
and the analyses of dissolved metals from samples
taken at the W4 sample site were recorded for
4 years after completion of the primary grouting
program in 1999.
Prior to the application of the source control
technology, flow results and visual observation
revealed that most of the flow into the workings
occurred in the Wl drift (Figure 6-2). After
grouting, flow in Wl was reduced; however, flow
recorded from the W2 and W3 drifts indicate that
they were not affected by the grout emplacement
even with the close proximity of the workings.
The last flow measurements taken inside the
Miller Mine workings were on October 8, 2001.
On that date, 53% of the flow was from the Wl
drift and 47% was from the combined flows from
the W2 and W3 drifts that were measured at the
W2 weir. As the flow at W4 increased from
1.9 gpm to approximately 3.0 gpm in September
2003, it is unknown where the increase in flow
originated because access to the underground mine
workings was restricted.
The water quality of the water discharging at the
Miller Mine portal (W4) improved as a result of
the application of the source control technology.
Water quality improvement was directly related to
the reduction of influx into the Wl drift. The
metal concentrations from Wl were diluted as the
water from Wl merged with water from the W2
drift. The concentration of dissolved metal in the
portal discharge significantly dropped after the
application of each phase of the source control
technology. Iron and Zn concentrations were
reduced the most when comparing 1999 pregrout
and 2003 postgrout analytical results. The
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dissolved concentrations for Pb did increase;
however, due to the decrease in flow, Pb loads
were on average reduced by 50%.
As calculated, the average percent reduction of
dissolved metal loads was greater than 50% for all
metals analyzed. Both Zn and Fe loading was
reduced as much as 90%, and the Phase II/Stage II
grouting each assisted with further reduction in the
metals loading (Figure 6-8). The most significant
reduction in metals loading at the Miller Mine
portal, 600 ft from the Wl drift, included:
- Fe from 3.05 Ib/d to 0.02 Ib/d;
- Mn from 0.46 Ib/d to 0.09 Ib/d; and
- Zn from 0.06 Ib/d to 0.004 Ib/d.
Most of the dissolved metals concentrations were
reduced by an order of magnitude or more. Not
only is this a direct result of the reduction of the
flow containing heavy metals, but it is also a result
of the encapsulation of the ore/sulfide zones by the
grout. The technology application eliminated the
possibility of AMD formation by eliminating
either the oxygen or water on the surface of the
sulfide material. Due to the emplacement of grout,
the water was unable to flow in the highly
fractured, sulfide zones; thus, water flows in areas
where the sulfide content is lower and does not
contain heavy metals. This reduces the amount of
metals loading in the Miller Mine discharge and as
flow increases to the maximum of 3 gpm, the
metals loading is low.
8.1 Lessons Learned
Future grout programs or any additional grouting
of the Miller Mine should consider the suggestions
listed below.
Obtain core samples from the site to determine
relative hydraulic conductivity of the fracture
pattern prior to beginning coring and drilling
operations. Occasionally core description and
intact core is available. This data may be
helpful in selecting an optimum grout for a
given fracture system.
Install an in-line mixer if expansive grouts are
used to thoroughly mix grout and water prior
to injection. This would ensure that all
injected grout would be activated and expand
within expected time constraints.
Install an in-line heating system to reduce the
viscosity of the grout during injection if a
grout of lower viscosity is unavailable. This
would reduce the viscosity of a viscous grout,
provide for greater hydraulic conductivity of
the fracture system, and increase the
likelihood that the fracture system would be
sealed.
Consider using the Jackleg drill system more
frequently. This would provide the
opportunity to drill and grout many short holes
in the vicinity of known seeps and conduits.
In addition, this drill system is more mobile
and can be used in tight and confined areas.
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9. References
1. U.S. Environmental Protection Agency,
Site Characterization and Materials
Testing Report: Underground Mine
Source Control Demonstration Project,
MWTP, Activity III, Project 8, MWTP-
145, April 1999.
2. U.S. Environmental Protection Agency,
Quality Assurance Project Plan
Underground Mine Source Control
Demonstration Project, MWTP, Activity
III, Projects, MWTP-112R1, October
1999.
3. Johnson, E.A., Geology and Gold
Deposits of the Confederate Gulch-White
Gulch Area, Broadwater County,
Montana, Montana College of Mineral
Science and Technology, May 1973.
4. U.S. Department of the Interior, Water and
Power Resources Service, Ground Water
Manual, pg. 257-285, 1981.
5. Houlsby, A.C., Construction and Design
of Cement Grouting, A Guide to Grouting
in Rock Foundations, A.S.T.C., F.I.E.
Aust, F. ASCE, John Wiley & Sons, Inc.,
New York, pg. 123-133, 1975.
6. Western Regional Climatic Center,
Historical Climatic Information, Monthly
Precipitation Listing, Monthly Totals,
1998 - 2001, http://www.wrcc.dri.edu/
CLIMATEDATA.html
7. U.S. Environmental Protection Agency,
USEPA Contract Laboratory Program
National Functional Guidelines for
Inorganics Data Review, 1994.
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