Christopher Mark

NIOSH, Pittsburgh, Pennsylvania, United States

thomas M. Barczak

NIOSH, Pittsburgh, Pennsylvania, United States

introDuCtion

The 20th century witnessed two revolutionary transformations in underground coal mining technology. 

At the beginning of the century, hand loading was 

nearly universal. Mechanical loading began the first 

great change in about 1930 (Figure 1), and this process was completed when continuous miners supplanted drilling and blasting during the ’50s and ’60s 

(Anon, no date). The rapid growth of longwall mining during the last two decades of the century was the 

second revolutionary development. 

Neither of these new mining methods could 

have succeeded without equally dramatic developments in ground control technology. If roof bolting 

had not replaced timbers, mechanized room-andpillar mining could not have reached its potential. 

Modern longwall mining is similarly inconceivable 

without heavy duty, self-advancing shield supports.

Clearly, these innovations in support technology have had an impact on mine safety as well. 

While roof falls once claimed hundreds of lives each 

year, today annual fatalities are usually numbered in 

the single digits (Pappas et al., 2000). The fatality 

rate has also improved, measured on either a per-ton 

or a per-hour basis (Figures 2 and 3). Yet a careful 

analysis of the history shows that improvements in 

safety did not occur in lock-step with technological 

developments. The safety culture, which encompasses regulatory mandates, company-specific 

safety policies, and other social and legal aspects, 

has also been central because it determines the level 

of risk that miners take (or are exposed to) while 

underground. 

GrounD ControL DurinG tHE 

HanD-LoaDinG Era

At the turn of the century, roof support was considered the responsibility of each individual miner. It 

was his duty to “examine his working place before 

beginning mining work, to take down all dangerous 

slate, and make it safe by properly timbering the area 

before commencing to mine coal.” The mine operator was responsible for delivering timbering materials to each working place, and the foreman checked 

that it was installed properly on his daily visit (Paul 

and Geyer, 1928).

The miners often had considerable discretion 

about the amount of support they installed. In weak 

shale roof, posts were set 2.5 to 5 ft apart (Paul et al., 

1930), but where the roof rock was strong, no posts 

might be set at all. An important element in early roof 

support systems were “safety posts,” which were set 

at the end of the track to protect miners while they 
loaded the coal or prepared the face for the next shot 

(Figure 4) (Paul et al., 1930). These temporary supports required extra time and effort, and their use was 

often at the discretion of the miner.

0

100,000

200,000

300,000

400,000

500,000

600,000

700,000

1900 1910 1920 1930 1940 1950 1960 1970 1980 1990

Underground Bitumn Producti0on (million tons)

Longwall Face

Continuous

Conventional

Hand Loaded

figure 1. underground coal mining methods of 

the 20th Century

0.00

0.20

0.40

0.60

0.80

1.00

1.20

1.40

1900 1910 1920 1930 1940 1950 1960 1970 1980 1990 2000

Years

Roof Fall Fatalities/Million Hrs 

figure 2. roof fall fatalities in underground 

coal mines

1

10

100

1000

1900 1920 1940 1960 1980 2000

Years

Roof Fall Fatalities 

(log scale)

figure 3. u.S. roof fall fatality rate for underground coal mines (fatalities per million 

employee hours)

figure 4. Hand loading with a safety post

Under these circumstances, it is understandable 

that much of the blame for roof fall accidents was 

placed on the inexperience and carelessness of the 

miners themselves (WVU Press, no date). In 1912, 

the USBM asked:

How can you, the miner, escape harm from 

roof falls?” The answer was: “Be careful...do not take the risk of loading a car 

before putting up a prop...set extra posts, 

even though they are in your way (Rice, 

1912).

Over time, however, safety professionals began to 

recognize that “a condition responsible for many 

fatalities from falls of roof is the absence of any 

policy on the part of management with respect to 

systematic methods of roof inspection and support” 

(Paul, 1927). Encouraging mine managers to prepare, promulgate, and enforce a systematic timbering 

plan became a key element in Bureau’s roof control 

efforts.

The USBM also exhorted miners to “comply 

with systematic methods of timbering, where such 

systems have been adopted, and exercise judgment 

in placing additional posts for your own protection.” 

But as long as the typical mine foreman was responsible for about 80 miners, and seldom spent more 

than 5 minutes with each one during a shift, enforcing timber plans would present a challenge.

tHE iMPaCt of MECHanizED MininG 

on GrounD ControL

Between 1930 and 1948, the portion of underground coal that was loaded by machine rose from 

less than 10 pct to nearly 2

⁄3 (Dix, 1988). In many 

ways, machine mining actually made roof support 

more difficult. Loading machines required a propfree front in which to work. The machine operator 

was usually protected by posts and crossbars, but 

the helper had to venture into the unsupported face 

zone (Figure 5) (Coal Age, 1945). A detailed and 

widely reported Bureau of Mines study conducted 

in 1951 concluded that “mechanical operations are, 
to a considerable degree, more dangerous from the 

standpoint of roof falls” then hand loading, “notwithstanding that much closer supervision is maintained in such operations.” The study also found that 

74% of roof fall fatalities occurred wit)hin 25 ft of 

the working face, and that 3 out of 4 of these took 

place by the last permanent support (between the last 

support and the face) (Forbes et al.,1951). The same 

study also concluded that:

Regardless of roof conditions, minimum 

standards of roof support suited to the 

conditions and mining system of each 

mine should be adopted and followed....

The judgment of the person should never 

be substituted for the minimum support 

required in the systematic roof support 

plan (Forbes et al.,1951). 

The difficulties posed by traditional timber supports increased as the early track-mounted loading 

machines were replaced by crawler-mounted ones. 

When rubber-tired shuttle cars that carried coal away 

from the face were introduced in the late 1930s, timbering became a critical bottleneck in the mechanical mining process. Little wonder, then, the Bureau’s 

Edward Thomas wrote in 1948 that: 

The more progressive mining companies 

are continually searching for improved 

types of roof support that will give maximum protection and at the same time offer 

minimum interference with the preparation 

and loading of coal (Thomas et al., 1948).

figure 5. Machine loading with timber support

tHE introDuCtion of roof BoLtinG

Even as Thomas wrote those words, the roof bolt was 

emerging as the leading candidate “temporary legless 

support” in machine mining (Thomas et al., 1948). As 

a support, roof bolts are superior to timbers because 

“timbers offer support after the strata they are supporting have failed; whereas roof bolts reinforce the 

roof rock, which contributes to its own support.” 

The St. Joe Lead Company was the first major 

mining company to make extensive use of roof bolts, 

beginning in the 1920s. Early in 1947, C. C. Conway, 

Chief Engineer for the Consolidation Coal Company 

in St. Louis, visited one of the St. Joe mines near 

Bonne Terre, Missouri, and was impressed with the 

roof bolts he saw there (Jamison, 2001). He determined to try them at Consol’s Mine No. 7 near 

Staunton IL. The first roof bolts were installed in 

Mine No. 7 using hand-held stopper drills (Figure 6). 

figure 6. installing roof bolts with a hand-held 

stopper drill

The anchors were expansion shells “similar to those 

used to support trolley wire,” though slot-and-wedge 

type anchors like the ones “ordinarily used in the 

metal mines” were also employed. 

The USBM was apparently involved in the roof 

bolt trials at Mine No. 7 almost from the beginning. 

Subsequently, the Bureau initiated a major effort 

to promote the use of roof bolts nationwide. They 

touted both the safety and efficiency advantages of 

the new supports (Thomas, 1951). Safety features 

included:

• A systematic support “within inches” of the 

working face

• Can’t be dislodged by blasting or equipment;

• Improved ventilation (because of less air 

resistance)

• Reduced accumulation of explosive coal dust 

(because places can be cleaned more thoroughly) (Thomas et al., 1949)

Economic advantages included:

• A reduction in the time required to load a place 

by 15–50%

• Potential for widening rooms
• Faster haulage

• Reduced labor cost for roof support and material handling

Since the Bureau was without regulatory powers, 

it had since 1910 “mastered the art of prodding 

operators into implementing new technologies that 

resulted from its scientific investigations (Aldrich, 

1995). The Bureau’s roof bolting campaign had to 

overcome numerous barriers to the new technology. 

The first was the cost and availability of the required 

equipment. The development of carbide alloy insert 

bits was essential because it made it possible to drill 

holes cheaply in hard rock (Thomas, 1956). Handheld stopper drills were already available at most 

mines, but they now required a mobile source of 

compressed air. Many of the first mines to install 

roof bolts built their own cars to carry the drills, 

compressor, bolts, and other supplies from face to 

face (Figure 7). 

figure 7. an early roof bolt machine

The bolts themselves were not readily available, and sometimes had to be fabricated in 

the mine’s own shop. Finally, miners were used to 

the reassuring presence of heavy timbers, and roof 

bolting seemed to be “reverse in principal to the 

old methods,” because it “appears at first glance to 

approximate holding oneself up by one’s bootstraps.” 

Between 1949 and 1955, numerous case histories of the successful application of roof bolting 

from all over the coalfields were reported in the 

mining press. For example, a major Bureau study 

in five southern West Virginia mines found that they 

“produced over two million tons of coal without a 

fatal accident and only 4 lost-time accidents, as compared with two fatal and 71 lost time accidents over 

a similar period when conventional timbering methods were used. Production increases ranged from 

0.86 to 10.7 tons per man shift.” That these results 

were achieved during the dangerous process of pillar 

recovery made them even more impressive. In northern WV, the success of roof bolting in pillar work 

allowed 7–10% more coal to be recovered (Flowers, 

1953).

Such figures led many safety professionals to concur with Joe Bierer of the West Virginia 

Department of Mines, that:

Herein lies a wonderful opportunity for 

the coal industry to bring about an epochal 

advance in safety for the mineworker, a 

humanitarian accomplishment to compare 

with the great social advances of recent 

years. No such immediately effective 

and readily confirmed benefit has derived 

from any other measure ever conceived, or 

devised, for safety in coal mines (Bierer, 

1952).

Unfortunately, even as roof bolting continued its 

dramatic expansion, overall fatality rates stubbornly 

refused to go down. The total number of roof fall 

fatalities declined, but three out five mining jobs 

also disappeared between 1948 and 1960. Each of 

the remaining miners actually had a greater chance 

of being killed in a roof fall than his counterpart in 

1948. 

The Bureau advanced two explanations for the 

frustrating lack of progress. The first was that when 

miners go beyond the last support, they are unprotected, regardless of what type of support is used. A 

1954 Bureau study had found that more than 50% of 

roof fall fatalities occurred in the unsupported space 

between the last row of supports and the face (Sall, 

1955). 

The second explanation for the lack of progress 

was that “many mines are now bolting where the 

method is marginal in the sense that perfect anchorage cannot be obtained.” Leon Kelly, a Bureau 

engineer in Vincennes IN, described this process in 

remarks he made to the 1950 Annual Meeting of the 

IL Mining Institute. He cited examples from three 

mines in his own experience in which the level of 

support had been reduced, and concluded:

In each case, when bolts were first used at 

the mine, everyone was more or less afraid 

of them and the pattern that was adopted 

was followed religiously. As time went on, 

and none of the bolts fell out, they were 

taken for granted and it was assumed that 

bolts would hold up the roof as long as 

there were bolts in the roof. Some operators are beginning to tell me that we are 

all overbolting, and naturally when they 

feel that way they will reduce the either 

the number of bolts or the length of the 
bolts they use.... If failures are accepted 

as a calculated risk, it is only a matter of 

time until a serious accident occurs (Kelly, 

1951).

Looking back, it seems clear that simply replacing 

timbers with bolts would not be sufficient to substantially reduce roof fall rates. The success of any support system in a particular application depends not 

just on the type of support, but also on the density 

of the pattern, the capacity of each unit, when they 

are installed, the quality of the installation, and many 

other factors including the span, the rock quality, and 

the ground stress.

Simply stated, roof bolts can only prevent roof 

falls if enough of them are installed. That costs time 

and money. The mine operator’s natural tendency 

was to adjust the expenditure to achieve an acceptable level of roof fall risk. Unless the mining culture 

was changed to reduce the acceptable level of risk, 

competitive pressures would mean that the new technology would be adapted to obtain the same results 

as before.

Moreover, it seems clear that roof bolting actually introduced a vicious new hazard, silica dust. The 

Bureau publicized research that showed dry drilling, with either pneumatic stoper or electric rotary 

drills, could result in silica dust concentrations up to 

200 times the recommended level of 5 million particles per cubic ft of air. Such concentrations were 

a serious menace to not just the drill operators, but 

also anyone working downwind in the return air. 

But in 1957, the Bureau reported that of the 424 

mines using roof bolts, just 8% employed water to 

allay dust, 35% employed dry dust collectors, and 

nearly half employed no means of dust control other 

than respirators (Figure 8). 

figure 8. Dry drilling using respirators for dust 

protection

There is little doubt that 

the prevalence of occupational lung disease among 

the generation of miners who worked in the dusty, 

mechanizing mines of the postwar period was higher 

than among most others in the past. Unfortunately, 

it seems that silica dust from roof bolting must have 

contributed to this terrible human toll.

tHE 1969 aCt

On Nov. 20, 1968, a massive gas and dust explosion 

destroyed the Farmington Mine and killed 78 miners. 

The disaster led directly to the Federal Coal Mine 

Health and Safety Act of 1969, which was signed 

into law by Richard Nixon. Under the Act, Federal 

inspectors were given much expanded enforcement 

powers, and a detailed set of health and safety standards was made mandatory for all mines (LewisBeck and Alford, 1980). The key ground control 

provision, one that Bureau ground control specialists 

had been advocating for two generations, was that 

each mine was required to develop a Roof Control 

Plan which included systematic support throughout 

the mine. Strict guidelines regarding bolt spacing, 

bolt length, entry width, and other ground control parameters were also included. Working under 

unsupported roof without safety posts was banned.

The results were quick and dramatic. In the six 

years following 1968, roof fall fatality rates plummeted by two-thirds, and they stayed approximately 

constant at that new level for the next 15 years. 

The improvement might have been due to regulatory enforcement, or to changes in safety standards 

implemented by the operators themselves, or to mandated safety technologies like protective canopies 

and Automated Temporary Roof Support systems. 

But there can be little doubt that the reduction was 

not due primarily to new technology, but rather to 

widespread application of technologies that were 

already available.

roof SuPPort anD LonGWaLL MininG

Modern longwall mining began in 1952 with a 

Bureau-sponsored trial at the Stotesbury mine of 

Eastern Gas and Fuel Associates near Beckley, 

WV (Haley and Quenon, 1954; Mason, 1976). This 

marked the beginning of mechanized extraction of 

a relatively large coal panel using a German coal 

planar. Roof control for the retreating longwall face 

was provided by wood cribs and mechanical 40-toncapacity friction props with I-beam caps. Three panels were successfully mined between 1952 and 1958 

producing an average of 530 tons per shift (Haley 

and Quenon, 1954).

The primary constraint to increased production from longwall operations during this era was 

roof support advance. Mechanical friction props and 
wood cribs were the predominant support system, 

and they had to be manually removed from roof contact, and in the case of the wood cribs, reconstructed 

as the face advanced. Support capacity during this 

era typically ranged from 25 to 75 tons per jack supplemented by 100 ton capacity wood cribs. The jacks 

were often damaged during the panel extraction and 

tended to be unstable in heavy strata conditions, 

which further reduced their effectiveness. 

The next innovation was the introduction of 

powered roof supports, which were hydraulically 

pressurized and advanced without manual labor. 

Eastern Associated Coal Company installed the first 

hydraulic self-setting and self-advancing roof support system in 1960 in a 52-in Pocahontas coal seam 

at their Keystone mine in southern West Virginia. 

These initial self-advancing roof supports were 

frame-type constructions where two hydraulic jacks 

were connected by a beam to form a frame construction as shown in Figure 9 (Chironis, 1977a). 

figure 9. Self-advancing frame-type roof support

The 

units were operated in sets of two with two frames 

linked together by a central shifting ram. As one 

frame remained set between the roof and floor, the 

other frame was lowered and advanced, thus creating 

a self advancing support system. Longwall installations using self-advancing frame supports were successful in strata conditions that provided easy caving 

of the roof behind the supports. However, there were 

also very serious failures when mines attempted to 

install these supports under more competent roof 

strata, such as massive sandstone and limestone 

geologies. 

The first high-capacity support system was 

installed in Island Creek’s Beatrice mine in 1966. 

This system featured 560-ton units that provided 

excellent control of the competent sandstone roof, 

and operated for a period of ten years without failure 

while producing a respectable 1,000 tons per shift of 

raw coal (Kuti, 1972). Nationwide, support capacity 

increased by a factor of 2 to 4 during the next few 

years, with four-leg supports providing resistances as 

high as 700 tons. 

An improvement in support design was also 

made with development of the chock support. The 

chock support can be thought of as a mobile crib. 

It essentially tied two frame supports together with 

a rigid canopy and semi-rigid base. This design 

improved structural stability and increased roof 

contact area over the previous frame-type systems. 

The increased roof cover helped reduce accidents 

caused by material falling from the immediate roof, 

especially during support advance. However, several 

failures were still reported due to the inadequacy of 

the chock support systems to operate in competent 

strata that caved in large pieces and exerted rotational moments or horizontal displacements to the 

support structure. 

It was not until the introduction of the shield 

support in 1975, with its improved lateral stability 

and more roof cover (Figure 10), that the last major 

hurdle to longwall utilization was overcome. The 

first shield-supported face installed in the United 

States was by Consolidation Coal Company (Consol) 

in 1975 at their Shoemaker mine near Moundsville, 

West Virginia. From the very beginning, encouraging 

results were obtained. The face produced 750,000 

tons in its first year of operation (Chironis, 1977b). 

The outstanding successes of these initial shield 

installations created a feverish quest for utilization 

of this technology. Shield utilization rose dramatically from their introduction in 1975 to 1982, when 

83 pct of the longwall faces were shield supported. 

By 1985, just 10 years after introduction, nearly 90 

pct of the operating longwall faces were supported 

by shields.

The basic shield design itself underwent several 

developmental changes during this era (Figure 11). 

The most significant design improvement was the 
incorporation of the lemniscate linkage into the caving shield assembly. Previous designs had the caving 

shield connected directly to the base, which caused 

the tip of the canopy to transcribe an arc as the shield 

was lowered and raised. The lemniscate system provided vertical travel of the canopy throughout its 

operating range, which minimized the unsupported 

span in front of the shield. 

figure 10. Comparison of chock and shield support constructions and stability

figure 11. Modifications of basic shield design

The USBM was extensively involved in promoting the utilization of longwall mining during the 

1970s. Bureau sponsored research totaled approximately 67 million dollars, which would be the 

equivalent of about $254 million today (BLS inflation calculator, http://data.bls.gov/cgi-bin/cpicalc.pl). 

Included in these efforts were: early demonstrations 

of the shield support system at York Canyon and Old 

Ben; demonstrations of thin seam longwall mining, 

single entry gateroad design, multi-lift extraction of a 

thick seam, steeply pitching seam extraction, and an 

advancing longwall trial. 

In 1979, the Bureau under contract to MTS 

Corporation built a sophisticated load frame at a 

cost of $6.8 million ($20.2 million in 2009) for conducting research on longwall support systems. The 

Mine Roof Simulator, as it is called (Figure 12), 

has the capability to provide controlled loading of 

up to 1,500 tons vertically and 800 tons horizontally 

over a 20-ft by 20-ft platen area to full scale support 

structures. This facility was instrumental in helping 

to identify design deficiencies in the early generation shield supports and to develop testing protocols 

taking advantage of the active loading capabilities of 

the simulator. Much of this effort centered around 

achieving proper load transfer through the caving 

shield lemnicsate assembly, particularly for horizontal shield loading (Barczak and Schwemmer, 1988a).
figure 12. Mine roof Simulator unique active 

loading frame used for full-scale support testing

As the shield continued to grow in capacity 

and the leg cylinders became increasing larger to 

accommodate the increase in capacity, two-leg shield 

designs became the standard in the industry. The 

advantages of the two-leg shield over the four-leg 

design is more efficient loading, less maintenance, 

and active horizontal loading whereby the inclination of the leg cylinders can impart through the 

canopy a confining force toward the longwall face 

(Barczak and Garson, 1986). This acts to restore the 

loss of confinement of the roof strata in the face area 

created by the formation of the gob as the immediate 

roof caves. The disadvantage of the two leg design 

is increased toe pressure that can create problems in 

soft floor conditions. This problem is mitigated by 

increased contact area and specialized base lifting 

devices when necessary. 

In addition to the increase in capacity, the 

shields have also increased in size. They have 

increased in length to accommodate one-web-back 

operations and larger face conveyors and deeper 

shearer webs, making these units over 20 feet in 

length by the end of the 20th century. The standard 

canopy width by the year 2000 was 1.75 m, replacing the 1.5 m designs that were prevalent during the 

1990s. The first 2.0-meter wide shield was installed 

in Foundation Coal’s Cumberland mine in 2003. 

This trend of increasing canopy width to 2.0 m is 

likely to continue. The 2.0-m width may represent an 

upper limit with current shield construction materials 

since wider shields weigh more, thereby requiring 

more effort to transport during face moves. 

The question of is bigger better or how much 

capacity is needed remains a debated issue as the largest shield capacity is now 1,750 tons. A consequence 

of the increase in capacity by increasing the leg cylinder diameter has been a proportional increase in 

shield stiffness (Barczak and Schwemmer,1988b). 

This results in increased loading per unit of convergence, so convergence which is beyond the direct 

control of the shield causes an increase in shield loading in proportion to the increase in shield stiffness. 

Less debatable has been the desire to maintain 

consistent and high setting forces. Modern electrohydraulic control technologies are capable of maintaining a set-to-yield ratio in the 0.6 to 0.8 range, 

provided there is sufficient flow capacity from the 

pumping system and the support hydraulics are well 

maintained. Unfortunately, the increase in shearer 

speed, automated shield advancement with multiple 

shields moving and simultaneous conveyor advancement, and wider longwall faces have placed increasing demands on the hydraulic power distribution 

system and consistent setting pressure remain a concern at many current longwall operations. 

DESiGn of LonGWaLL GatE Entry 

SyStEMS

On December 19, 1984, 28 miners entered the longwall section at the Wilberg Mine near Huntington, 

UT. They hoped to break the world’s record for oneday production from a longwall. Instead, a fire broke 

out at an air compressor located near the mouth of 

the section, and deadly smoke filled the headgate 

entries. Only one miner escaped.

During the subsequent investigation many 

safety experts insisted that miners died, not because 

there was a fire, but because they had no escape 

route. The miners had been forced to try to escape 

through the headgate because the tailgate from the 

Fifth Right longwall had been blocked by roof falls 

(Cocke, no date). As a result, new regulations were 

introduced that required longwall mines to maintain 

safe travel ways on the tailgate side of longwalls 

(30CFR75, 2000).

At the time, there was very little theory available to assist mine planners in sizing longwall pillars. 

Traditional pillar design methods were not appropriate because the long and narrow geometry of typical 

gate entry layouts makes a classic squeeze highly 

unlikely. In addition, the abutment loads applied to 

longwall pillars are significantly greater and more 

complex than the tributary area loading assumed by 

traditional pillar design methods. Effective longwall 

pillar design requires some way to estimate the abutment loads during all phases of longwall mining, 

from development all the way into the tailgate. 

Most importantly, however, the stability of the 

tailgate is not purely a pillar design problem. Studies 

conducted as early as the 1960s had concluded that 

“whether or not the stress [from an extracted longwall panel] will influence a roadway depends more 
on the strength of the rocks which surround the roadway itself than on the width of the intervening pillar” 

(Carr and Wilson, 1982). 

The USBM developed the Analysis of Longwall 

Pillar Stability (ALPS) method to answer these 

needs. First, design estimates of abutment loads were 

obtained through an extensive series of underground 

stress measurements. The greater challenge was how 

to integrate pillar design into a broader design method 

that focused directly on tailgate performance. 

The solution was to employ a sophisticated 

empirical technique. The Bureau conducted studies 

at more than half of all the longwalls in operation in 

the U.S., and documented historical gate entry performance at all of them. A new rock mass classification system, the Coal Mine Roof Rating (CMRR), 

provided a quantitative measure of the structural 

competence of the roof (Molinda and Mark, 1994). 

Multivariate statistical analysis provided simple, 

quantitative guidelines for sizing longwall pillars, 

and showed that when the roof is strong, smaller pillars can safely be used. 

The Bureau conducted an extensive program 

to disseminate the ALPS technology to the industry. 

ALPS was implemented as a user-friendly computer 

software package, which was widely distributed free 

of charge. A number of hands-on computer training 

sessions were held throughout the coalfields. Since 

the mid-1990s, ALPS has been considered the industry standard for longwall gate entry design.

In addition to this pillar design effort, an extensive development in standing support to provide 

increased stability in the tailgate entries occurred 

in the early 1990s. Timber supports in the form of 

wood cribs and posts were all that was used prior 

to mid-1980, when trials of concrete supports were 

common. Then, the 1990s saw a revolution in new 

support technologies, providing a wide range of support capabilities. Figure 13 (figure shows a distribution of several types of tailgate supports since this 

time. 

Wood Cribs

Timber Posts

Concrete Cribs

Eng Timber Cribs

Cables 

Yld Timber Props

Burrell Can

3-C Support

Steel supports

Pumpable cribs

1993

1996

2001

2005

0

10

20

30

40

50

60

70

80

90

SUPPORT USAGE, pct

figure 13. Distribution of tailgate support since 

1993 showing change in support practices from 

conventional wood cribs and posts into a variety 

of alternative support technologies

The soft response of the wood cribs that allowed 

too much roof movement to occur under abutment 

loading led a search to develop stiffer and higher 

capacity supports for longwall tailgate applications. 

However, a valuable lesson was learned from concrete 

supports that had a compressive strength and material modulus an order of magnitude higher than the 

conventional wood cribs. Any standing support cannot control the behavior of the overburden. Plain and 

simple, you cannot prevent all of the closure of a mine 

opening with any practical and economical support 

system, and if the support cannot survive this uncontrollable convergence, it will fail prematurely and 

not provide adequate roof support. These super high 

capacity and stiff concrete cribs were crushed in several longwall tailgates (Figure 14) (Barczak, 2006). 

figure 14. failure of high capacity concrete 

crib in longwall tailgate

This lesson has transpired into a goal to utilize the ground reaction concept for support design 

(Figure 15). The basic concept is to match the performance characteristics to the ground response, 

whereby the convergence developed in the longwall 

tailgate is proportional to the amount or capacity of 

support used. Numerical models are used to develop 
the ground response curves which are calibrated with 

in mine measurements of support loading and convergence. The onset of onset of strain-soften rock 

response, indicating that the rock has been loaded 

beyond its elastic range and is now yielding causing damage to the rock mass, defines the support 

threshold. Providing more support than this provides 

little benefit since the support cannot impact the elastic ground response and will produce a negligible 

reduction in tailgate convergence. Conversely, too 

little support will result in greater convergence and 

increased risk of instability leading to roof fall. 

Convergence

Support Pressure

A. Initial Stress Level 

C. Nadir

D. 

B. Yielding 

Convergence

Support Pressure

A. Initial Stress Level 

C. Nadir

D. 

B. Yielding 

Added support 

pressure

Risk of 

instability

Support Design Threshold

Little Benefit

Damaged Rock

Elastic Response

Yield Response

figure 15. Ground reaction curve for support design

The performance characteristics of the various 

roof support products are determined from full-scale 

testing in the NIOSH Mine Roof Simulator through 

a rigorous testing protocol that determines the limitations of the support system. This data is incorporated 

in the NIOSH Support Technology Optimization 

Program (STOP) which facilitates selecting the most 

effective support for a particular ground condition 

(Barczak, 2006 and 2008). 

NIOSH subsequently employed the same basic 

empirical approach to develop solutions for several 

other complex ground control problems. The software packages, particularly the Analysis of Retreat 

Mining Pillar Stability (ARMPS) and Analysis of 

Multiple Seam Stability (AMSS), have been widely 

adopted by the mining community. Each is based 

on extensive data bases of real world case histories, 

gathered from in-mine studies. Because they are so 

firmly linked to reality, they have met the mining 

community’s needs for reliable design techniques 

that can be used and understood by the non-specialist. 

Unfortunately, the 2007 Crandall Canyon 

mine disaster provided an unwelcome reminder of 

the importance of pillar design. The MSHA report 

on the disaster pointed to the “flawed” mine design 

as the root cause, because “the stress level encountered exceeded the strength of a pillar or group of 

pillars near the pillar line.” The report documented 

how the two pillar design software packages used to 

develop the mine plan, ARMPS and LaModel, had 

been employed improperly. Following the disaster, 

MSHA distributed a series of memorandums and 

other documents that strongly encouraged mine 

planners to use ARMPS in the pillar design process. 

The memorandums state that any design that does 

not meet the NIOSH criteria should be considered 

“complex and/or non-typical,” and be subject to 

more extensive MSHA review.

rECEnt aCCoMPLiSHMEntS 

anD futurE CHaLLEnGES

Fatalities inby supports: Fifty years ago, miners 

working inby supports were the victim in half of 

all underground roof fall fatalities. The practice of 

going inby supports was banned by the 1969 Act, 

but through the 1990s such accidents still accounted 

for nearly half of all roof fall fatalities. During the 

past six years, however, there have been just been 

two inby incidents. The progress is associated with 

the industry-wide educational campaign under the 

slogan “Inby is Out!” Another factor contributing to 

the reduction in inby fatalities is the development of 

the concept of the “Red Zone.” In many roof control 

plans, miners must now stay outby the second row 

of bolts, particularly when making extended cuts or 

turning a crosscut.
Pillar recovery: Progress has also been made 

in pillar recovery. Historically, retreat mining has 

been less safe than other underground mining techniques. Although the intentional caving that occurs 

is an unavoidable part of the retreat mining process, 

premature caving can cause roof falls that put miners 

at risk. A NIOSH study published in 2003 found that 

retreat mining elevated a miner’s risk of being killed 

in a roof fall by a factor of three. 

The same study also found that many of the 

fatalities had occurred where the mine was actually 

following its Roof Control Plan. As a result, NIOSH, 

together with MSHA, advocated that retreat mines 

increase the roof support they employ during retreat 

mining by leaving final stumps, using Mobile Roof 

Supports, and installing longer and stronger roof 

bolts. The widespread adoption of such safer retreat 

mining techniques seems to have made a difference. 

During the past four years there has been just one 

fatal roof fall, compared to an average of two per 

year during the previous decade. 

Rock and rib falls: More than 400 miners continue to be injured each year by rock falling from 

between supports, and 100 more are injured by rib 

falls. Together, these two categories also account 

for a large percentage of recent ground fall fatalities. Available technologies such as roof screen, rib 

bolting, and inside control roof bolters could dramatically reduce injury and fatality rates if they were 

used more widely. Further advances in these areas 

will likely be the next big advance in ground control 

safety.

Disclaimer

The findings and conclusions in this paper have not 

been formally disseminated by the National Institute 

for Occupational Safety and Health and should not 

be construed to represent any agency determination 

or policy.

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