Table I. Extended Cuts at U.S. Coal Mines

State Coal Seam Entry width (ft) CMRR Cut Depth (R) | Cover (®) Cut Status*
AL Blue Creek 205 70 30 7000 1
AL Blue Creek 20.3 30 25 7800 2
AL Mary Lee 211 36 33 300 I
L Herrin No. 6 155 38 30 650 1
IL Herrin No. 6 18 7 40 150 1
i Herrin No. 6 24 86 40 350 1
L Herrin No. 6 16.5 38 35 650 2
L Herrin No. 6 18 45 40 150 3
L Herrin No, 6 16 33 20 330 3
L Springfield No. 5 19.5 30 40 600 1
iN ‘Springfield No. 5 188 39 40 250 1
iN Springfield No. 5 188 29 15 250 3
KY Cedar Grove 18.5 30 40 600 1
KY Harlan 18 6 40 1800 1
KY Kentucky No. 11 172 a 200 2
KY Kentucky No. 11 183 33 14 7000 3
KY Kentucky No. 9 175 7) 40 150 1
KY Kentucky No. 9 19.1 38 30 350 1
KY Kentucky No. 9 20 38 36 350 2
KY Kentucky No. 9 175 35 20 150 3
KY Kentucky No. 9 18 35 17 1000 3
KY Pond Fork 19 35 30 300 1
KY Pond Fork 18 45 30 900 3
OH Freeport (L) 19.5 68 40 600 1
OH Freeport (L) 19.5 4 20 600 3
PA Kittanning (L) 17 47 40 500 1
PA Kittanning (L) 20 30 30 600 1
PA Kittanning (L) 17 45 40 300 2
PA Kittanning (L) 14 33 15 600 3
PA Pittsburgh 16.5 40 30 650 2
PA Pittsburgh 17 40 30 150 2
PA Sewickley 18.5 44 30 800 2
UT “De 20 35 1000 2
UT Hiawatha 20 70 40 1500 1
wv Coalburg 18.5 2B 40 300 1
wy Coalburg 20 76 40 300 1
wv Dorothy 18.5 39 40 500 1
WV Eagle 177 36 37 1000 1
wy Eagle 20.1 68 40 800 1
wv Eagle 20.1 56 40 800 2
wv Eagle 17.7 35 10 1000 3
wy Pocahontas No. 3 18 30 25 1200 2
wv Powellton 18.5 34 7 400 3
WY Hanna 183 B 40 700 1

‘Extended Cut Status: 1=Always Stable 2=Sometimes Stable 3=Never Stable


roof conditions were encountered in different areas within the
mine, resulting in a total of 44 case histories. The CMRR was
calculated from underground observations in each case.
Usually, the entry width was determined as the mean of at least
10 underground measurements, but in others it was the nominal
width supplied by the mine. The typical cut is the permitted
deep cut, except where reduced by "conditions." ‘The depth of
cover is typical of the area studied.

Mine officials were asked to rate how often they were able
to achieve a full extended cut, Responses like"always" and
"almost every time" were grouped in class 1 (“Always Stable").
Class 2 ("Sometimes Stable") included responses like "about
half the time." Cases where it was "rarely" possible to
complete an extended cut, or where company policy limited cut
depths when certain conditions were encountered, were
grouped in Class 3 ("Never Stable"). In one case, the roof was
reported to collapse as soon as the box cut was mined, reducing
the effective roof span to 14 ft. The "Never Stable" class also
included two mines that had not applied for an extended cut
permit because they felt their roof was too weak. Several other
mines that had not applied for other reasons, such as methane
control, were not included in the table,

‘The data is shown in table 1. All 8 cases where the CMRR_
was less than 36 fell into the "Never Stable" class, Only 2
"Never Stable" cases had a CMRR greater than 36. Of the 12
"Sometimes Stable" cases, 8 occurred where the CMRR was

less than 46. Where the CMRR was greater than 56, every case
was “Always Stable."

The multi-variate statistical technique of logistic regression
was employed to quantify other relationships within the data.
Logistic regression is used where the outcome variable is
"dichotomous" (has two levels). When asked to discriminate
between the "Never Stable" group and the other two the
only significant variable was the CMRR. For CMRR>38, only
2 immediate collapses are misclassified.

The results were more enlightening when the "Never
Stable” and "Sometimes Stable" groups were combined and
compared with the "Always Stable" group. Now both the
CMRR and the depth of cover (H) were identified as
statistically significant. An overall accuracy of 87% (6
misclassifications) was achieved with the equation:

CMRR = 40.9 + (H/100) @

Figure 4 plots the CMRR against the depth of cover, along
with equation (2). The most likely explanation for trend
towards less stable extended cuts as the cover increases is the
greater stress level. The vertical stress increases in proportion
with the depth, and studies have shown that the horizontal
stress typically increases twice as rapidly as the vertical (Mark
and Mucho, 1994).

Depth of Cover (m)
0 100 200 300 400 500 600
100
90
80
70
© 60
s 50
° 40
30 ° © Never Stable
20 4 Sometimes Stable
10 @ Always Stable
sees Extended Cut Equation
0
0 200 400 600 800 1000 1200 1400 1600 1800 2000

Depth of Cover (ft)

Figure 4. Relationship Between CMRR and Depth of Cover for Extended Cut Data Set.


Entry Width (m)

4 45 5.0
80 a 7 7
@ Never Stable
70 A Sometimes Stable
HI Always Stable
604“ Extended Cut Equation
w —— Bieniawski's Equation
£6

La om

13 14 15 16

18 19 20 21 22

Entry Width (ft)

Figure 5, Relationship Between CMRR and Entry Width for Extended Cut Data Set.

Figure 5 plots entry width against the CMRR for the data
set. The statistics indicated that entry width was less
significant than depth of cover in predicting the performance of
extended cuts, The best equation with CMRR and the entry
width (W,) was equation 3 (85% overall accuracy, 7
misclassifications):

CMRR = 19.2 + 1.64 W, @)

Note that this equation is very similar to the one calculated
from Bieniawski's data (equation 1), which is also plotted on
figure 5. There is a difference of about 6 CMRR points in the
intercept because equation (3) separates the "Sometimes
Stable” from the “Always Stable” groups, while equation (1)
corresponds to “always immediate collapse." Most of the
“Never Stable" group falls below the Bieniawski line, as
predicted,

The best equation that combines the effects of both depth
and span was:

CMRR = 18.6 + (H/100) + (1.2 W, ) (4)

CONCLUSIONS

The study confirmed that the stability of extended cuts is
determined primarily by the roof quality, as measured by the
CMRR. The entry width and the depth of cover are secondary
factors.

‘The results provide some simple guidelines for predicting
the performance of extended cuts, When the CMRR is less
than 38, it is unlikely that extended cuts will be feasible.
When the CMRR is above 55, extended cuts should be routine.
For intermediate roof conditions, extended cuts are more likely
to be troublesome as the roof span widens or as the cover
deepens. Equation (4) can be used to predict the when
problems are likely to be encountered.

REFERENCES

Beerkircher M. D. (1994). Monterey Coal Company's
Longwall Project. Proc. IL Mining Institute., Collinsville, 1
pp. 85-93.


APPLICATION OF THE COAL MINE ROOF
RATING (CMRR) TO EXTENDED CUTS

by
Christopher Mark

National Institute For Occupational Safety And Health
Pittsburgh Research Laboratory, Pittsburgh, Pa 15236


Bieniawski, Z. T. (1989). Engineering Rock Mass
Classifications. Wiley, NY, 251 pp.

DeMarco, M. J. (1984), Yielding Pillar Gate Road Design
Considerations for Longwall Mining. Paper in New
Technology for Longwall Ground Control: Proceedings of the
USBM Technology Transfer Seminar, USBM SP 94-01,
pp. 19-36.

Ferm, J. C., and G. A. Weisenfluh (1981). Cored Rocks of
the Southern Appalachian Coalfields. USBM Contract
No. H0230028, Library of Congress #81-51290, 93 pp.

Grau I, R. Hand E, R. Bauer (1997). Ground Control
Worker Safety During Extended Cut Mining. Proc. 16th Intl
Conference on Ground Control in Mining, Morgantown, WV,
pp. 283-288.

Harwood, C.,M. Karmis, C. Haycocks, and J. Luo (1996).
Optimizing Secondary Tailgate Support Selection. Proc. 15th
Intl. Conf, on Ground Control in Mining, Golden, CO, pp. 469-
476.

Luo, J., C. Haycocks, and M. Karmis (1997). Gate Road
Design in Overlying Multiple Seam Mines. SME Preprint
97-107, 11 pp.

Mark, C, and F. E. Chase (1994). Design of Longwall Gate
Entry Systems Using Roof Classification, Paper in New
Technology for Longwall Ground Control: Proceedings of the
USBM Technology Transfer Seminar, USBM SP 94-01,
pp. 5-18,

Mark, C., G. M. Molinda, A. P. Schissler , and W. J. Wuest
(1994), Evaluating Roof Control in Underground Coal Mines
Using the Coal Mine Roof Rating. Paper in the Proceedings of
the 13th Conference on Ground Control in Mining, pp. 252-
260.

Mark, C. and G, M, Molinda (1996). Rating Coal Mine
Roof Strength from Exploratory Drill Core. Proceedings of the
15th International Conference on Ground Control, Golden, CO.,
pp. 415-428.

Mark, C. and T. P. Mucho (1994). Paper in New
Technology for Longwall Ground Control: Proceedings of the
USBM Technology Transfer Seminar, USBM SP 94-01.

Molinda, G. M. and C. Mark (1994). The Coal Mine Roof
Rating (CMRR)—A Practical Rock Mass Classification for Coal
Mines. USBM IC 9387, 83 pp.

Molinda, G. M. and C. Mark (1996). Rating the Strength of
Coal Mine Roof Rocks. USBM IC 9444, 36 pp.


ABSTRACT

Since it was first introduced, the Coal Mine Roof Rating
(CMRR) has been widely accepted as a tool for geologic
characterization and mine planning. This paper discusses the
application of the CMRR to another practical ground control
problem.

Extended cuts (cuts greater than 6 m (20 ft) in length) are
commonly used with remote control continuous miners.
Extended cuts can greatly increase productivity, but they have
been associated with a number of fatal roof fall accidents.
‘When extended cuts are attempted in weak roof, the roof may
collapse before it can be bolted, causing hazardous conditions,
Until now, it has not been possible to predict where conditions
may not be suitable for extended cuts. In this study, data on the
CMRR and extended cut experience were collected at 36 mines
in 7 states. It was found that when the CMRR was greater than
55, deep cuts were routine in nearly every case. When the
CMRR was less than 37, extended cuts were almost never
taken. Between 38 and 55, extended cuts were feasible
sometimes but not others. The data also shows that extended
cuts are less likely to be stable if either the entry span or the
depth of cover increases.

INTRODUCTION

On July 15, 1997, four coal miners were killed by roof falls
in three separate incidents. This tragic coincidence was a
dramatic reminder that roof falls remain the single greatest
hazard faced by underground coal miners.

One reason that roof falls have proved so stubborn a
problem is that mines are not built of man-made materials like
steel or concrete, but rather of rock just as nature made it. The
structural integrity of coal mine roof is greatly affected by
natural weaknesses including cracks, small faults, and layering.

To make matters more difficult, the geologic processes that
formed it varied in space and time, so engineering properties of
the roof can change dramatically from mine to mine (and even
within individual mines!).

Engineers have had difficulty obtaining quantitative data on
the strength of rock masses for design. Traditional geologic
reports contain valuable descriptive information, but seldom
include engineering properties. On the other hand, laboratory

strength test results are inadequate because the strength of
small rock samples are only indirectly related to the strength of
the rock mass.

‘To help quantify the engineering properties of mine roof,
the Coal Mine Roof Rating (CMRR) was proposed (Molinda
and Mark, 1994). The CMRR combined 20 years of research
on geologic hazards in mining with worldwide experience with
rock mass classification systems. Field data was collected from
nearly 100 mines in every major U.S. coalfield. Cost-sharing
cooperative research agreements were signed with the Cyprus,
Ziegler, and Peabody coal companies to support the research.

‘The CMRR weighs the geotechnical factors that determine
roof competence, and combines them into a single rating on a

scale from (to 100. The CMRR makes four significant
contributions:

‘© Focuses on the characteristics of bedding planes,
slickensides, and other discontinuities that weaken the
fabric of coal measure rock;

© Applies to all U.S. coalfields, and allows meaningful
comparison even where lithologies are quite different;

© Concentrates on the ability of the immediate roof to
form a stable structure, and;

© Provides a methodology for geotechnical data
collection.

Originally, the data for the CMRR was collected at
underground exposures like roof falls and overcasts. To make
it more generally useful, procedures were developed for
determining the CMRR from drill core (Mark and Molinda,
1996). In addition, more than 2000 point load strength tests
were conducted on samples of common coal measure rocks
(Molinda and Mark, 1996). The samples were classified
according to the popular pictorial rock core numbering system
developed by John Ferm (Ferm et al., 1981).

‘The CMRR has found numerous applications in ground
control design. A study conducted at 44 longwall mines found
that tailgate performance was largely determined by the CMRR
and the ALPS pillar stability factor (Mark and Chase, 1994).
Guidelines for sizing longwall pillars based on the CMRR have
since been widely implemented. The same study found
significant correlations between the CMRR and both entry
width (figure 1) and the intensity of roof support. Another
study determined that yielding pillar gate entry designs have
only been successful when the CMRR was greater than 50 and
the pillar’s width-to-height ratio was less than 5 (DeMarco,
1994).

‘Coal Mine Roof Rating

(© Satisfactory cases a Jnsatisfactory cases

Figure 1. Relationship Between CMRR and Width at U.S.
Longwall Mines (after Mark and Chase, 1994).

Data have also been presented that relate the incidence of
roof falls to the CMRR and intersection span (figure 2). These
were based on observations at five underground mines (Mark et
al,, 1994). The CMRR has lately been incorporated into
guidelines for multiple seam mine design (Luo et al., 1997),


hazard analysis and mapping (Wuest et al., 1996),
support selection (Harwood et al., 1996), and feasibility studies
(Beerkircher, 1994). The Mine Safety and Health
‘Administration has used the CMRR in fatal accident
investigations, and at least three major coal companies have
recently taken steps to integrate the CMRR into their
exploration programs.

COMRR = 35 5

Roof falts (A)

COMRR = 45 4
L_~ omar = 65

ga

Roof falls per 3 Km,

%e & & A BG
Intersection span (m)

Figure 2. Predicted Incidence of Roof Falls as a Function of
Intersection Span (Sum-of-the-diagonals) and CMRR (Mark et
al., 195).

STABILITY OF EXTENDED CUTS

Extended (deep) cut mining is where the continuous miner
advances the face more than 20 ft beyond the last row of
permanent supports, The development of remote control, spray
fan systems, and flooded bed scrubbers provided the technology
to enable continuous miners to take deep cuts and still comply
with mining regulations. Since 1989 the number of mines with
approvals for extended cuts has increased from 206 to 399
(Grau and Bauer, 1997). About 75% of all underground man
hours are worked at mines with extended cut permits.

In practice, many mines with permits only take extended
cuts when conditions allow them. Where the roof is competent,
extended cuts are routine. AL the other extreme, when the roof,
is very poor miners may not even be able to complete a 20 ft
cut before the roof collapses. Since mining personnel should
never be beyond roof supports, falls of unbolted roof should not
bea major hazard. Yet between 1988 and 1995, extended cuts
may have been a factor in 26% of all roof fall fatalities in
underground coal mines (Grau and Bauer, 1997), The ability to
identify areas where extended cuts might collapse prematurely
could be very useful to mine planners and regulators.

An extended cut in a coal mine is actually a special case of
‘an unsupported span. The stand-up time of unsupported spans
is one the fundamental issues in rock engineering. The basic
relationship that governs stand-up time was originally
formulated by Austrian tunneling engineers (Bieniawski,
1989):

© Fora given rock mass, a tunnel’s stand-up time
decreases as the roof span becomes wider, and;

© Fora given roof span, a tunnel's stand-up time
decreases as the rock mass quality becomes poorer.

Using data collected from numerous tunnels and mines,
Bieniawski (1989) was able to quantify this relationship

(figure 3). Bieniawski used the Rock Mass Rating (RMR) as
the measure of rock quality. His data indicates that an
unsupported 14 ft wide tunnel would be expected to collapse
immediately if the roofs RMR was less than 33. If the tunnel
was 20 ft wide, immediate collapse would be expected if the
RMR was less than 41. Equation (1) expresses the relationship
for this range of tunnel spans (approximately the range
encountered in underground coal mining):

RMR=13+1.4W, oo)

Where W, is the entry width (ft).

Meee | Hore ere | Menine | Your
€
i: Seat &
3 Fite
Ae

VE

+ o

‘Stand-up Time, Hours

Figure 3. Relationship Between RMR, Tunnel Span, and
Stand-up Time (Bieniawski, 1989).

Because roof bolting normally takes place within several
hours of mining, the collapse of an extended cut may be
considered "immediate."

‘The CMRR was developed to represent the unique
characteristics of layered coal measure rock, while the RMR is
‘more appropriate for jointed hard rock. However, the CMRR
‘was designed to be equivalent to the RMR, so that the
CMRR/unsupported span/stand-up time relationship should be
nearly the same in both systems (Molinda and Mark, 1994).

The fundamental relationship between rock quality, span,
and stability was apparent in two of the studies reported earlier.
As figure | shows, longwall mines with poor quality roof have
“naturally" gravitated to narrower entries. Similarly, figure 2
shows that the incidence of roof falls increases either as the
span increases or as the CMRR decreases. Extended cuts
provide an opportunity to evaluate the stand-up time of coal
mine roof without the influence of roof bolts.

RESEARCH RESULTS

The stability of extended cuts was investigated at 36 mines
in 7 states between 1994 and 1997, In eight mines, different