International Journal of Rock Mechanics & Mining Sciences 44 (2007) 210–221

Short communication

Case history of the response of a longwall entry subjected to

concentrated horizontal stress

C. Marka,, W. Galeb, D. Oylera, J. Chenc

a

NIOSH, Rock Safety Engineering Branch, Pittsburgh Research Laboratory, P.O. Box 18070, Pittsburgh, PA 15236, USA b

Strata Control Technologies, Wollongong, NSW, Australia c

RAG Pennsylvania, Waynesburg, PA, USA

Accepted 29 June 2006

Available online 25 September 2006

Abstract

The National Institute for Occupational Safety and Health (NIOSH) Pittsburgh Research Laboratory (PRL), RAG Pennsylvania and

Strata Control Technologies of Australia collaborated in an intensive study of ground behavior, reinforcement performance, and stress

redistribution at the Emerald Mine in Southwestern Pennsylvania. The study site was a longwall tailgate subjected to a severe horizontal

stress concentration. Field measurements indicated that the stresses applied to the study site nearly doubled during longwall mining,

resulting in roof deformations extending to a height of 4.8 m (16 ft) above the entry. A computer simulation of the field site was

conducted using FLAC-2D, incorporating a broad range of rock behaviors and failure mechanisms. Comparison between the

measurements and the simulation showed that the model was able to capture the most significant aspects of the roof and support system

behavior, particularly, the extensive slip along bedding that created a partially destressed ‘‘softened’’ zone in the immediate roof. The

model also showed that supplementing the normal roof bolt support pattern with cable bolts would allow the entry to survive a further

20–25% increase in the applied horizontal stress. Such information could have very practical application to the design of roof support

systems for coal mines.

Published by Elsevier Ltd.

1. Introduction

Nearly 1500 roof falls occur each year in US underground coal mines, creating serious safety hazards and

operational impacts. The cost of the support installed to

prevent roof falls approaches $1 billion annually. Despite

more than 50 years of research, no roof support design

methodology has gained wide acceptance [1].

During the past 25 years, it has become evident that

horizontal stresses are a significant contributing factor in

many of the roof falls that occur underground [2]. An

important breakthrough was the recognition that the

stresses observed in mines are caused by global platetectonic forces [3]. Stress measurements have confirmed

that the horizontal stress is typically two to three times the

vertical in mines in the eastern US. Moreover, the longwall

mining process can concentrate the horizontal stresses in

specific, critical locations [4,5].

Stress concentrations that affect longwall gate entries are

of particular concern (Fig. 1). These develop because the

stress cannot be transmitted through the caved gob areas

behind the longwall. Roof falls in either the tailgate or

headgate are major safety hazards because they can disrupt

ventilation or block vital travelways. They can also cause

expensive production interruptions.

Most horizontal stress concentrations can be avoided

through proper longwall panel orientation and sequencing.

In some instances, however, the geometry of the reserve or

some other consideration may limit mine planning options.

For example, where one longwall panel is significantly

longer than an adjacent, previously mined, panel, the

resulting ‘‘stress window’’ can be very difficult to control

(Fig. 2). Under these circumstances, the mine must prepare

for severe ground conditions.

In the US, a common response to poor ground

conditions is to install supplemental supports, such as

ARTICLE IN PRESS

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1365-1609/$ - see front matter Published by Elsevier Ltd.

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Corresponding author. Tel.: +1 412 386 6522; fax: +1 412 386 6891.

E-mail address: cmark@cdc.gov (C. Mark).
cable bolts. The interaction between supplemental supports

and primary roof bolts has received relatively little research

attention, however [6]. The traditional way to study roof

support performance has been with field measurements.

During the 1990s, measurements of roof deformation and

roof bolt load were made at a number of sites in the US

[7,8]. These studies provided numerous insights into factors

that effect bolt performance, including installed tension,

bolt capacity, and grout annulus. Unfortunately, none of

these studies was combined with stress measurements, or

with extensive geologic characterization and rock mechanics testing. Moreover, sophisticated numerical modeling like that employed in some other countries has had

relatively limited application to mining ground control in

the US [9–11].

The study described in this paper was the result of a

collaboration between the National Institute for Occupational Safety and Health (NIOSH), RAG Pennsylvania

(now Foundation Coal) and SCT Operations of Australia.

Its goals were to:

measure the stress changes and ground deformations

resulting from an anticipated longwall stress concentration;

evaluate the performance of typical US primary

supports (roof bolts) in response to high horizontal

stress;

investigate the interaction between supplemental support systems (cable bolts) and primary supports; and

refine the numerical modeling techniques required for

analyzing support alternatives for high stress coal mine

applications.

2. Study site

The site of the study was the tailgate of the 11 North

longwall at the Emerald mine, located in Greene County,

Pennsylvania (Fig. 3). The site was chosen because it was

anticipated that the extension of 11 North beyond the start

line of 10 North would result in a significant horizontal

stress concentration [5]. From past experience at Emerald

Mine, a horizontal stress window like the one created by 11

North could be expected to cause severe loading to be

applied to the crosscut and tailgate entry. The location thus

provided a rare opportunity to study the roof failure

process as the applied horizontal stress increased during

the progression from development through longwall

mining.

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Fig. 1. Concentrations of horizontal stress caused by longwall panel

extraction.

Fig. 2. A ‘‘stress window’’ created when horizontal stress is funneled

between a pre-existing gob area and an approaching longwall panel.

Fig. 3. Emerald mine and study site.

C. Mark et al. / International Journal of Rock Mechanics & Mining Sciences 44 (2007) 210–221
The Emerald Mine operates in the Pittsburgh coal

bed in Southwest Pennsylvania, cutting roadways approximately 2.1–2.4 m (7–8 ft) high and 4.9 m (16 ft) wide.

Approximately 0.3 m (1 ft) of roof shale is cut from the

roof of the roadways. A geologic column of the mine roof

obtained from a vertical core hole drilled at the site is

shown in Fig. 4. The roof may be roughly divided into

three units:

a sequence of coals and weak, slickensided shales in the

lowest 2.7 m (9 ft);

a slightly stronger claystone sequence from 2.7 to 5.4 m

(9–18 ft); and

a significantly stronger limestone above 5.4 m (18 ft).

The in-mine coreholes barely reached the limestone, so

its thickness and strength were estimated from a nearby

surface corehole. The low uniaxial compressive strength

and RQD for the bolted horizon results in an estimated

coal mine roof rating (CMRR) of 37 [12].

Rock properties for the models were obtained from a

combination of previous testing at the mine and from tests

of underground core samples obtained from the study site.

Of particular importance were angled core that were drilled

at the study site and subjected to multi-stage triaxial

testing1 in order to determine bedding plane strength. The

bedding plane strength test results are presented in Fig. 5

for the two shale units. The UCS profile and bedding

cohesion profile used in the model is presented together

with the laboratory core test results in Figs. 6 and 7,

ARTICLE IN PRESS

UCS 17.7 MPa

0.0

0.5

1.0

1.5

2.0

2.5

3.0

3.5

4.0

4.5

Redstone

Limestone

Claystone

Gray

Shale

Claystone/

Fireclay

Bone

Rider

Coal

RQD 0%

Coal

Depth, m

5.0

5.5

Fireclay

w/thin UCS 24.4 MPa

Ls, slips,

& clay

veins

Black

shale

w/coal streaks

RQD 40%

RQD 36%

UCS 36.4 MPa

RQD 50%

RQD 0%

RQD 89%

UCS 31.3 MPa

RQD 27%

RQD 29%

Fig. 4. Composite core log from the study site, showing rock physical

properties. Data from the vertical corehole drilled in the study crosscut

and from a surface corehole.

Fig. 5. Bedding plane test results from angled core.

Fig. 6. UCS profile used in the model together with rock property data

derived from testing.

1

The multi-stage triaxial test aims to define the rock strength envelope

for a number of confining pressures (2, 5, and 10 MPa) using a single

sample. A stiff, servo controlled testing machine must be used in the

deformation control mode. The sample is loaded progressively until the

onset of initial fracture, and then the confining pressure is increased to the

next stage and the test continues. At 10MPa confining pressure the sample

is taken to full failure and then the confining pressure is released slowly

and the residual strength monitored to determine the post failure strength

envelope. When angled core samples are tested, the failure takes place

along a bedding plane, and the applied stresses are resolved into the shear

and normal stresses as shown in Fig. 3. An early application of multi-stage

triaxial testing is described in Dolinar et al. [13].

C. Mark et al. / International Journal of Rock Mechanics & Mining Sciences 44 (2007) 210–221
respectively. The layer-to-layer variations in the model

profiles were derived from both the test results and

inspection of the roof core. The in situ strength of coal

within the model was set at 6.5 MPa (900 psi), which is

typical of the bulk strength of coal.

The stress field has been measured in several locations in

the Emerald and the adjacent Cumberland mine, but not at

the specific study site. These measurements were reviewed

to assess the range of stresses anticipated at the test site.

The major stress was oriented N701E. The regional stress

field appears to be relatively consistent with a lateral

tectonic strain [14] of approximately 550 microstrain. This

means that the maximum stress is inferred to be

approximately 11 MPa (1600 psi) for a rock having a

Young’s Modulus of 20 GPa (3 million psi). The horizontal

stress within other rock units will be different and

dependent on their elastic properties. The minor horizontal

stress is estimated to be approximately half of the major

stress. Vertical stress is approximately 5 MPa (700 psi) and

related to overburden of approximately 200 m (650 ft).

Primary roof support used in the area of this study

consisted of three 22 mm (7

8 in) diameter, 2.4 m (8 ft) long

combination bolts. The bolts were installed with 1.2 m (4 ft)

resin cartridges in 35 mm (1–3

8 in) boreholes. The yield load

of the bolts was 19 ton (21 ton) with an ultimate capacity of

28 ton (31 ton). At one of the instrumentation arrays,

supplemental support consisting of rows of three cable

bolts were installed between the rows of primary bolts. The

cable bolts were 3.6 m (12 ft) long, 15 mm (0.6 in) diameter,

and they were partially grouted with 1.2 m (4 ft) of resin.

3. Field instrumentation

At the site, two monitoring arrays were installed in a

crosscut and a third in the adjacent tailgate entry (Fig. 8).

The two crosscut sites, labeled sites B and C, are the focus

of this paper. Conditions in these sites were more severe

than in the tailgate site because the crosscut was oriented

less favorably relative to the regional maximum horizontal

stress.

Instruments were deployed to measure roof movements,

support loads, and roof stress changes. Fig. 9 is a cross

section showing the instrumentation in each array. The

most detailed roof movement data were provided by multipoint sonic extensometers, with magnetic anchors located

approximately every 0.3 m (1 ft) to a height of 5.8 m

(19 ft) above the roof line. These were supplemented by

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Fig. 7. Bedding plane cohesion profile used in the model together with

rock property data.

H2 H3

H7

H5

H4

N

Pumpable Cribs

Surface

Corehole

(overburden 198 m)

Vertical

Corehole

HI Cells

B array

C array

11 North Panel

Stopping

-5-10-15-20-25 0 5 10 15 20 25 30

H6

H1

Fig. 8. Site map. Distances shown are meters from the center of the

crosscut.

Coal Coal

0 ft

0 2 m

Strain gauged bolts

Sonic

Extensometer

Mechanical

Extensometer

5

Fig. 9. Typical instrument array. The C site included nine rows of three

cable bolts, and one row instrumented with U-cells in addition to the

instruments shown.

C. Mark et al. / International Journal of Rock Mechanics & Mining Sciences 44 (2007) 210–221
mechanical, three point extensometers that were monitored

remotely.

Loads on the lower, 1.2 m (4 ft) long ungrouted portions

of the combination roof bolts were measured using strain

gauges installed inside the bolts using the technique

developed by Dr. Hani Mitri at McGill University [15].

The instrumented bolts were obtained from Emerald Mine

and prepared at McGill. Holes were drilled into the head of

each bolt to accept a single strain gauge, near the bolt head,

and electrical connectors. The bolts were individually

calibrated to 13 ton (14 ton) by NIOSH and found to have

a highly linear load-to-strain-gauge-signal response in the

elastic range. NIOSH also tested combination bolts to

determine the post-yield load-deformation relationship, so

that estimates could be made of the bolt loads beyond

yield. During the study the strain gauged bolts performed

well, with no failures. Hydraulic U-cells were used to

monitor the loads on three cable bolts at site C, but

unfortunately only one provided useful measurements.

Horizontal stress changes were measured by seven

CSIRO hollow inclusion (HI) cells that were installed in

the roof above the gateroad pillar inby the crosscut site.

Each HI cell consisted of an array of 12 strain gauges

arranged to allow three-dimensional determination of

stress changes. The cells were grouted in place using an

epoxy designed specifically for HI cells to provide coupling

between the rock and the gauges in the cell. Seven cells

were installed with a 100% success rate using procedures

developed by Dr. Jan Nemcik of SCT. Before installation,

the holes were undercored to identify appropriate setting

zones and to obtain samples for physical property testing.

The stress cells were installed in a fan pattern as shown

in Fig. 10. Cell H7 was installed towards the top of the

lower shale unit, while cell H6 was placed in the bottom of

the limestone. The other five cells were installed to provide

information on the stress distribution within the upper

claystone unit. The strain data obtained from the HI cells

were reduced using software developed by the US Bureau

of Mines [16]. For presentation purposes, the stress

changes have been transformed into principal stresses

perpendicular to the crosscut.

4. Modeling approach and model used

Detailed monitoring studies conducted in coal mines in a

number of countries have shown that the mechanisms of

failure about excavations can be highly complex, involving

fracture of rock, failure of bedding or joints, buckling of

parted rock, and slip along weak surfaces [1,4,9]. The use of

computer simulation therefore requires a detailed geotechnical characterization of the strata and stressfield, and must

incorporate the many potential failure mechanisms [17,18].

Generalization of the rock properties or the rock mass

section on the basis of averaged properties has been found

to limit the capability to reproduce actual rock mass

behavior.

The code used in the model is FLAC which has been

modified to employ rock failure routines developed by SCT

Operations. The constitutive models used by SCT are very

similar to the strain-softening, ubiquitous joint model (SU)

included with the latest release of FLAC2D [19]. Rock

failure is based on Mohr–Coulomb criteria relevant to the

confining conditions within the ground. A broad range of

potential failure modes are simulated including:

shear fracture of intact rock;

tension fracture of the rock;

bedding plane shear; and

tension fracture of bedding (bedding separation).

The stability of pre-existing jointing, faults or cleat is

also addressed in the simulations where appropriate. The

model simulates new or re-activated rock fracture and

stores the orientation of the fractures.

In the SCT constitutive model, as well as Itasca’s SU, the

intact rock matrix exhibits strain-softening post-failure

behavior. A weakness plane of any orientation is also

included in the model, and this weakness plane can also

exhibit strain-softening behavior. This constitutive model is

most appropriate for coal measure rocks where the intact

rock is strain-softening and one dominant weakness plane

exists, namely horizontal bedding. In both SCT’s and

Itasca’s constitutive models, cohesion, friction angle,

dilation angle and tensile strength are specified as functions

of the ‘‘plastic strain.’’ Depending on the nature of these

functions, a variety of complex hardening and softening

behaviors can be produced.

The in situ strength of the rock materials is reduced to

0.58 of the laboratory unconfined compressive strength

(UCS). This lab-to-field scaling factor was originally was

suggested by [20], and is routinely used in all SCT model

studies.

The model geometry is presented in Fig. 11 with the

various rock layers characterized by their laboratory UCS.

The typical element size in the region of interest is

approximately 20  10 cm (8  4 in). This element size is

ARTICLE IN PRESS

H1 H2 H3 H4

H5

H6

H7

HI Cell Locations

Scale, m

Vertical Corehole

Angled Corehole

0 2 Study crosscut

Outby rib

Section view N77.5°E

Inby rib

Fig. 10. Section view of HI cell locations.

C. Mark et al. / International Journal of Rock Mechanics & Mining Sciences 44 (2007) 210–221
fine enough to capture geologic variations that may be

important at the coal mine entry scale.

Rock bolts were included in the model. The bolts were

bonded in the upper 1.2 m with a free length to the roof

line. Yield of bolts was 19 ton (21 ton) with a pullout force

of approximately 15 ton/ft. Cable bolts were also modeled

as partially bonded with a yield capacity of 25 ton (27 ton).

Care was taken to simulate the shear strength and stiffness

properties of the resin/rock interfaces of the bolts. All of

the bolt properties were adjusted to account for the row

spacing in the 2-D model.

The modeling sequence followed the actual entry

development process in that the entry was first excavated,

then the outside roof bolts were installed, and finally the

center bolt was installed. The model was allowed to

respond to each of these steps. If cable bolts were used in

a model, they were placed last.

For this model, a plane of symmetry was used at the

roadway center line. The use of symmetry speeds the

modeling process particularly in this case where a range of

stress conditions or support systems is to be evaluated.

5. Results: field measurements

The study site was subjected to increasing horizontal

stress during each of the three mining phases:

the initial entry development;

the start-up of the 10 North longwall; and

the retreat of the 11 North longwall past the study site.

As the study site was developed, evidence of stress

damage included a small ‘‘cutter’’ along the inby rib of the

crosscut. Ground conditions at the study site notably

deteriorated as the 10 North longwall began mining, even

though the start-up room was located approximately 100 m

(300 ft) away (Fig. 3). This was attributed to development

of the horizontal stress concentration at the corner of the

10 North gob. Roof mapping conducted at the time

showed that roof failure had extended some distance from

the start-up room, and included at least one large roof fall.

Apparently the breakup of the roof near the start up room

caused a gradual transfer of stress to the vicinity of the

study site.

The third phase, the approach and mine-by of the 11

North longwall, resulted in further roof deterioration

associated with the stress concentration at the tailgate

corner. Conditions were not as severe as had been

encountered in past stress windows at the Emerald Mine,

however, and the study site did remain accessible almost

until the face passed.

Stress measurements: The HI cells were not installed until

well after the entries were developed, so the stress changes

associated with entry development were not recorded.

Following the start-up of the 10 North longwall, the two

HI cells that were in place showed roof stress increases on

the order of 3.5 MPa (500 psi). The rate of stress increase

gradually moderated, but never completely stabilized (see

Fig. 12).

Fig. 13 is a snapshot of the stress changes that had been

measured by the end of Phase 2 (startup of the 10 North

longwall). The major principal stresses are all compressive,

and range in magnitude up to 20 MPa (3000 psi). The

orientations of the principal stress increases imply that the

immediate roof of the crosscut yielded or ‘‘softened,’’ and

was not capable of transmitting additional horizontal

stress. As a result, the stress was forced higher into the roof

above the crosscut. The extensometer data that will be

discussed later provides further evidence of the roof

softening.

After the start-up of the 11 North longwall, definitive

stress changes began to be observed in all HI cells when the

face was approximately 192 m (640 ft) inby the center of the

study tailgate intersection. After the face reached 27–32 m

(90–110 ft) inby the cells the rate of stress change greatly

accelerated. In most cases the stresses increased as the face

approached, typically by about 7 MPa (1000 psi). The

direction of the maximum principal stress increases

continued to be in a generally sub-horizontal orientation,

directed around the softened roof of the crosscut (Fig. 14).

One notable exception was H7, installed in the lower shale

unit, where the stresses remained almost constant as the

face approached.

As the face passed by the site, nearly every HI

cell measured an immediate horizontal stress reduction

of about 10 MPa (1400 psi), indicating that caving

of the immediate roof above the panel resulted in stress

relief.

In summary, it appears that the mining-induced stress

concentration approximately doubled the original in situ

stress. The orientations of the stress increases measured by

the HI-cells indicates that the stresses predominantly

passed over, and not through, the softened (yielded) roof

immediately above the crosscut.

Roof deformation: Initial extensometer readings were

first made several days after development, so the data do

not include the initial roof sag, but by mid-November the

roof at all three sites had stabilized with less than 5 mm

(0.2 in) total deformation (Fig. 15). However, the maximum height of roof movement was about 3.3 m (11 ft),

well above the top of the bolting horizon at 2.4 m (8 ft).

ARTICLE IN PRESS

Fig. 11. Model geometry and UCS of rock units.

C. Mark et al. / International Journal of Rock Mechanics & Mining Sciences 44 (2007) 210–221
During Phase 2, the 10 North start-up period, the

maximum height of roof deformation at the B and C arrays

increased to 3.9–4.2 m (13–14 ft.), with total deformations

of 16 and 10 mm (0.6 and 0.4 in), respectively (Fig. 15).

Further evidence of roof softening during this period is

apparent in the measurements of bolt loads and roof

movements. Maximum roof strains in excess of 1.5%

were measured at two points in the B extensometer, while

the C extensometer had one location where the strain

exceeded 1%.

As the 11 North face approached, both of the sonic

extensometers in the crosscut recorded significant roof

movements. Total deformations in each increased by about

20 mm (0.8 in) and the height of roof movement reached

4.9 m (16 ft). Most of this deformation occurred during the

last 23 m (75 ft) of advance before the face reached the

crosscut.

While the two arrays differed in the timing and

magnitude of the roof deformation, it was significant that

the deformation process followed a broadly similar pattern

in both locations. The maximum height of movement was

approximately 5 m (16 ft), and significant roof strains

occurred both near the top of the combination bolts (at

2.4 m) and approximately 1.5 m (5 ft) above them.

ARTICLE IN PRESS

Estimated

softened roof

zone

0

Scale, m

10

Stress scale, MPa

0

Study Crosscut

H7

H1 H2

H3

H4

H5

H6

2

Fig. 13. Principal stress changes after the completion of the 10 North

longwall and prior to mining the 11 North longwall in the plane

perpendicular to the study crosscut, with respect to initial HI cell readings.

Stress changes with arrows are tensile or stress relief; stress changes

without arrows are compressive.

-18

-16

-14

-12

-10

-8

-6

-4

-2

0

2

11/30 12/28 1/25 2/22 3/22 4/19 5/17 6/14 7/12

Stress change, MPa (- is compressive)

Normal Stress ⊥ to Face

Normal Stress || to Face

Vertical Stress

Stress changes with respect to 12/10/2002 initial reading.

130 94 63 33 331511 -9 -31 -56

7/13 7/14 7/15 7/16 7/17 7/18 7/19 7/20 7/21 7/22 7/23 7/24

-2611

-2211

-1811

-1411

-1011

-611

-211

189

Stress change, psi

Face distance from study crosscut intersection, m

Time scale change

11 North startup

971

Fig. 12. HI cell stress changes from cell H4 located 4.0 m above the roof and 7.6 m into the gateroad pillar. Vertical and normal stresses parallel and

perpendicular to the face are shown. The shear stresses are not shown. Negative values are compressive changes.

Estimated

softened roof zone

0

Scale, m

10

Stress scale, MPa

0

Study Crosscut

H7

H1

H2 H3

H4

H5

H6

2

Fig. 14. Principal stress changes through 0330 on July 21, 2003 in the

plane perpendicular to the study crosscut, with respect to initial HI cell

readings. The face is 2 m (8 ft) outby the study crosscut tailgate entry

intersection. Stress changes with arrows are tensile or stress relief; stress

changes without arrows are compressive.

C. Mark et al. / International Journal of Rock Mechanics & Mining Sciences 44 (2007) 210–221
Roof bolt loads: Fig. 16 shows the average bolt loads that

were measured during the course of the study at the two

instrument arrays. When they were first installed, the initial

roof bolt loads were typically 3–8 tonnes. During the next

several weeks the loads increased, particularly at the B-site,

the site without cable bolts. In fact, Fig. 16 shows that

while the initial loads were higher at the C-array, the loads

on the B-array bolts soon exceeded those on the C-bolts

and then remained higher throughout the course of the

study. The lower loads on the C-bolts were attributed to

the presence of the supplementary cable bolts (although the

ventilation stopping near the C array may also have

provided some support [21]).

With the start-up of the 10-North longwall, the loads on

the B-array bolts increased another 3–5 to 12–22 tonnes,

with the highest readings indicating that bolts were

reaching yield. At the C array bolt load increases were

somewhat less, to 8–15 ton.

By the time the 11 North face was within 40 m (130 ft) of

the site, all of the B-array bolts were in yield. One C array

bolt went into yield suddenly when the face was 105 m

(340 ft) inby, and the center bolt yielded almost as the face

passed. The cable bolts reached their maximum recorded

loads at this time, apparently below yield. After the face

passed the study crosscut the mechanical extensometers

indicated that roof deformations large enough to load a

cable bolt well into yield (more than 50 mm) took place

below the tops of the cable bolts. Bolt loads at the crosscut

arrays reached a maximum shortly after the face had

passed, and then began to decrease.

Unfortunately, the loading record for the cable bolts was

incomplete due to instrument problems. However, the

nearly 50 mm (2 in) of measured roof movement, along

with the lower bolt loads measured at array C, both

support the indication from the measurements that are

ARTICLE IN PRESS

0.0

0.5

1.0

1.5

2.0

2.5

3.0

3.5

4.0

4.5

12/10/2002

1/08/2003

4/29/2003

7/15 (face: +143)

7/17 (face: +90)

7/18 (face: +66)

7/19 (face: +39)

7/20 (face: +26)

Sonic Extensometer

(B array)

Depth, m

5.0

5.5

Displacement, mm

20100.0 30 5040

Top of cable bolts

Bolt tops

Fig. 15. Profile of roof movement in the study crosscut at the B array

sonic extensometer, from December 2002 through the last reading on July

20, 2003, with roof lithology. Face distances shown are in meters, with the

face in by the extensometers. Depths are referenced to the roof line.

Laboratory yield of combination bolts was 19 tonnes, ultimate load 28

tonnes. Sensor readings over 19 tonnes (yield) have been estimated 

by assuming a post yield stiffness 1/15 the elastic stiffness. 

C Array average

B Array average

Averaged Bolt Loads, tonnes

11/210/5 11/30

10

5

0

12/28 2/221/25 5/174/193/22

25

20

15

7/126/14 7/13 7/167/157/14 7/17 7/24

130

Face Distance from study crosscut, m

7/21

11

7/207/197/18

1533336394

7/237/22

-56-31-9

Time scale change

971

11 North

startup

Fig. 16. Bolt loads. Strain gauged bolts only, from the B and C arrays. Each curve is averaged from three bolts.

C. Mark et al. / International Journal of Rock Mechanics & Mining Sciences 44 (2007) 210–221
available that the cables did assume a significant portion of

the roof load. Recent field and numerical model studies

[6,22] both indicate that supplemental supports probably

reduced the loads that would have otherwise have been

applied to the roof bolts.

6. Stress path used in the model

The far-field (boundary condition) horizontal and

vertical stresses applied to the model at each successive

stage of the modeling process is the ‘‘stress path.’’ The

stress path represents the incrementally increasing stresses

that developed during the initial development of the study

crosscut and the subsequent extraction of the two longwall

panels. The stress path employed in this study is presented

in Fig. 17. It was estimated from the anticipated in situ

stresses, the stress change measurements made at the site,

and measurements made at similar sites in past studies.

Another indication of the ‘‘far field’’ stress increase

associated with longwall mining at this site was the

approximately 10 MPa (1400 psi) of stress relief that was

measured following the passage of the longwall face.

7. Modeling results: roof movement

The comparison between the model results and the field

measurements relies heavily on the roof extensometer data.

The approach has been to compare the extensometer

results at approximately equivalent total displacements at

the roof line. In this way, the nature and style of

deformation within the roof which causes the total

displacement can be compared. This allows the performance of different support systems to be evaluated using

the same criteria that are employed in the mine—namely,

how much roof movement is taking place. Therefore, it is

not necessary to employ surrogates such as the maximum

stress or a safety factor.

Fig. 18 compares the roof deformations calculated in the

model to those measured underground. Model results from

three stress levels on the model loading path are shown.

Also shown are a number of deformation profiles measured

at the two crosscut sites. The figure indicates that the model

was able to capture the deformation pattern common to

both sites, showing a very close correlation in both

deformation style and height of movement. It should also

be noted that the model indicated that some deformation

occurred even above the top of the extensometers.

Fig. 19 illustrates the location and nature of the rock

failure processes that are associated with the roof

deformation profiles. Upon initial development (Fig. 19a)

bedding plane shear occurs readily within the section and

early in the deformation process. As stress levels increase

shear fracture of the weaker shale units occurs. Higher

stress levels cause shear of the stronger materials together

with additional bedding plane shear. However, total roof

deformations are relatively minor until the applied far-field

stress reaches approximately 15–16 MPa (2200 psi;

Fig. 19b). Beyond that stress level, significant roof

deformation develops and progresses higher into the roof

section (Fig. 19c).

8. Modeling results: roof bolt forces

The roof bolt forces developed in the model and roof

bolt load data from sites B and C are presented in Fig. 20.

To make the comparison, the average bolt loads were

determined for three levels of roof displacement. In

general, there is a fairly wide range in average bolt load

at the monitoring sites, however, the overall forces

developed in the model are consistent with the range as

monitored. This provides an indication that the bolt-strata

interaction is being simulated in a realistic manner

consistent with the site response.

The results indicate that under the in situ development

stress state (less than approximately 11–12 MPa (1600 psi))

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Fig. 17. Stress path modeled representing increasing stress during development and longwall extraction operations.

Fig. 18. Comparison of modeled and measured roof displacement

profiles.

C. Mark et al. / International Journal of Rock Mechanics & Mining Sciences 44 (2007) 210–221
roof conditions would be anticipated to be good and well

controlled by the bolt pattern placed. The roof bolts

continue to be well under yield load up to a stress level of

approximately 15–16 MPa (2200 psi). Once significant roof

deformations begin to occur, however, the bolt loads

rapidly increase.

The model results (Fig. 20) show that the addition of

cable bolts initially has little effect on bolt loads. However,

once major roof movements begin to occur, the cables

assume enough load to delay the onset of yield in the roof

bolts. The cable bolts appear to develop their full capacity

after the roof bolts yield.

9. Modeling results: stress redistribution about the cross cut

caused by roadway deformation

The measurements made during the study showed that

the additional stresses were redirected above the immediate

roof of the crosscut even before significant roof deformations had occurred (Fig. 13). The model indicated that the

bedding plane shear that developed early in the deformation process could be sufficient to cause significant stress

redirection even though the displacement and visual

deformation of the roadway was low. An example is

presented in Fig. 21 for an equivalent displacement of

approximately 20 mm (8 in).

The amount of horizontal stress transferred within the

initial 2 m (7 ft) of roof at the center of the roadway is an

indication of the stability of the roof section and the

requirement for reinforcement of the rock. When the

bolted roof section has lost its integrity and become

‘‘softened,’’ the horizontal stress it can carry is reduced.

Once extensive rock fracturing causes roof softening, the

reinforcing action of the reinforcement is the primary

design task.

In summary, it appears that the model results are

consistent with the monitored and observed behavior

of the roadway available at the site. This provides

confidence that the model is simulating the rock deformation processes influencing roadway stability and reinforcement interaction. This confidence allows for realistic

assessment of various bolt patterns within the stress

conditions anticipated at the site, as discussed in the next

section.

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Fig. 19. Roof deformation in the model at three stress levels:

(a) 11–12MPa, (b) 16 MPa and (c) 19MPa.

Fig. 20. Roof bolt forces developed relative to horizontal stress.

C. Mark et al. / International Journal of Rock Mechanics & Mining Sciences 44 (2007) 210–221
10. Application of the modeling results to reinforcement

design

In this section, the effect of hypothetical bolt patterns

were evaluated to demonstrate the influence of reinforcement patterns on roadway behavior.

The bolt patterns assessed in this example were:

1. 3, 1.8 m (6 ft), long fully grouted bolts.

2. 3, 2.4 m (8 ft), long combination bolts.

3. 3, 2.4 m (8 ft), long combination bolts with 2, 4 m (13 ft)

cables.

Fig. 22 shows the roof displacement within the roof

section. Each bolt system has a characteristic ‘‘limit’’ at

which point the roof deformations begin to rapidly

increase. It is significant that the deformation does not

follow a smooth curve towards failure, but rather abruptly

goes from ‘‘controlled’’ to ‘‘uncontrolled’’ movement. This

seems to conform with underground experience, where

‘‘good’’ conditions often seem to ‘‘suddenly’’ go bad.

For the standard 2.4 m (8 ft) bolt pattern, the roof

maintains integrity up to approximately 15–16 MPa

(2200 psi), which is within the range anticipated during

development at high angles to the regional stress field. The

addition of supplemental cable bolts would allow the roof

to cope with an additional 3–4 MPa (500 psi).

On the other hand, 1.8 m (6 ft) bolts reach their limit at

just 12 MPa (1700 psi), which could change a situation of

probable roof control success during development in the

cross cut direction to one of potentially difficult roof

conditions. These results are consistent with expectations,

and indicate that various combinations of reinforcement

can significantly modify the deformation limit of the strata

section.

11. Conclusions

The study confirmed that the ‘‘stress window’’ that

resulted from the mining geometry did cause a dramatic

increase in the horizontal stress at the site. The measurements indicated that the stress approximately doubled as

the mining progressed.

In the vicinity of the entry, these stresses were

significantly reoriented by the deformation and failure that

developed in the mine roof. The zone of measured ‘‘roof

softening’’ extended above the height of the bolts, causing

the additional horizontal stress to be transferred higher

into the roof. The softened zone did not become a fully

‘‘detached block’’, however. Even site B, which was

supported by just the normal roof bolt pattern, did not

collapse. However, the roof bolts at site B did yield much

earlier in the study than did those at site C, where

supplemental cable bolts had been installed. It appeared

that the cable bolts assumed a portion of the loads that

would otherwise have been applied to the roof bolts.

The numerical model was able to capture the most

significant aspects of the behavior of the study site. The

modeled roof deformations matched the measured ones in

both magnitude and location. A critical element in the

model’s success was its ability to simulate the extensive

bedding plane slip that developed in the roof.

Additional modeling showed that each support pattern

had a critical level of applied horizontal stress beyond

which roadway conditions dramatically deteriorated. For

example, 1.8 m (6 ft) bolts perform adequately up to an

applied stress level of approximately 12 MPa (1700 psi),

and increasing the length of the roof bolts to 2.4 m (8 ft)

allows the entry to survive stresses up to 15–16 MPa

(2200 psi). Adding cable bolts would increase the capacity

of the support system by approximately another 20–25%.

It seems, therefore, that the model could be used to select

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Fig. 22. Displacement of the roof section for various bolt patterns.

Fig. 21. Stress redirection about the roadway caused by localized rock

failure and subsequent changes in bulk material properties during

roadway development.

C. Mark et al. / International Journal of Rock Mechanics & Mining Sciences 44 (2007) 210–221
the optimum support pattern if the expected stress increase

could be estimated.

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