208

ABSTRACT 

The objective of this is paper is to review the current state of 

knowledge and practice in highwall mining (HWM). HWM has 

become a widely-applied method in surface mining, commonly used 

alone or in conjunction with contour or slot mining. It provides 800-

feet to 1,200-feet of additional recovery when the economic stripping 

ratio is reached in contour mining or in slot mining when surface 

access to a reserve is limited. A significant attribute of the highwall 

miner is its versatility. HWM has been used successfully to mine 

 abandoned pre-reclamation law highwalls, 

 points or ridges uneconomic to mine by underground or 

other surface methods, 

 outcrop barriers left adjacent to underground mines, 

 separate benches of the same seam where the parting 

thickness or quality differences between benches render 

complete extraction uneconomic, 

 previously augered areas containing otherwise inaccessible 

additional reserves and 

 close or widely spaced multiple seams. 

The theory and design methods to assess roof, pillar, and floor 

stability are presented followed by three case histories. Simple design 

charts for sizing HWM web and barrier pillars are also presented. A 

recommended web pillar width may be obtained from the design 

charts given the overburden depth, the HWM cut width, and the 

mining height. Given the depth and panel width for a set of HWM 

cuts, another set of charts gives a suggested barrier pillar width. 

The case histories, from Northern and Southern Appalachia are 

used to illustrate the application of rock mechanics to quantify the 

stability of the highwall, roof, web pillars, and floor. The case 

histories involve 1) mining through a previously augered highwall, 2) 

mining under back-stacked spoil and 3) selective mining of closely 

spaced benches of the same seam. 

Because each site is unique, the appropriate pre-mining 

geotechnical analyses range from the calculation of roof, web pillar, 

and floor bearing capacity stability factors to detailed numerical 

modeling of the auger and underground mine workings. When 

operating in the vicinity of existing underground mine or auger 

workings, the determination of ground deformation and strains 

resulting from highwall mining is a necessary facet of a ground 

control investigation. 

INTRODUCTION 

Auger and highwall mining has evolved from a secondary 

production method to become an integral part of many surface 

mining operations throughout the U.S. and the world. Auger 

mining in the Appalachians began sometime in the mid 1940s when 

surface miners turned vertical blast hole drills horizontal at the 

economic limit of a surface pit to recover more coal from the 

highwall. Modern auger and highwall mining has evolved from 

this humble beginning into a highly productive, sophisticated, hightechnology, surface coal mining technique requiring four to five 

miners per shift. 

Volkwein, et al. (1) review the evolution of auger and 

highwall mining systems in the U.S. including the earliest augers 

dating from the mid 1940s, early highwall mining concepts such as 

the “Carbide Miner1

”, the “Push-button Miner” (1970s), the “Edna 

Miner” and the Metec miner (1980s) and finally several continuous 

haulage concepts in the 1990s including CONSOL’s “Tramveyor” 

and Arch Coal’s “Archveyor.” Arrowsmith (2) and Fiscor (3) 

discuss recent developments in highwall mining techniques. 

Currently, two manufacturers dominate the market for highwall 

mining systems with each having about 30 systems in operation. 

The Superior Highwall Mining Company (SHM) (4) developed the 

Superior Highwall Miner, based upon the Metec design. Mining 

Technologies Inc., now International Coal Group, Inc. – ADDCAR 

Systems, LLC (5) developed the ADDCAR system. 

The two systems are similar in that a cutter head, consisting of 

a continuous miner head and gathering arms is attached to a series 

of coal transport modules that are added as the cutter head is 

advanced into the coal seam. The machines are operated from an 

enclosed, climate controlled cab, located at the rear of the unit. The 

SHM, shown in photograph 1 uses 20-foot long rectangular “push 

beams” that are dropped in place as the cutter head is advanced. 

Twin spiral augers are fully enclosed within the “push beams” and 

are used to move the coal away from the face. 

The ADDCAR system relies upon a series of conveyor cars 

that utilize conveyor belts to transport coal from the face, as shown 

in photograph 2. Another significant difference between the two 

HWM systems is the guidance and monitoring systems utilized to 

 

1

Mention of specific product or trade names does not imply endorsement by 

NIOSH. 

Analysis of Highwall Mining Stability - The Effect of Multiple Seams 

and Prior Auger Mining on Design

David Newman, President

Appalachian Mining and Engineering, Inc. 

Lexington, KY 

R. Karl Zipf, Mining Engineer 

NIOSH-Pittsburgh Research Laboratory 

Pittsburgh, PA 
209

direct and monitor the HWM while cutting to reduce out-of-seam 

dilution and to maintain a constant web width. The guidance and 

monitoring system technology, which is outside of the scope of this 

paper, uses a combination of sophisticated PLC control, video 

cameras, and methane monitors, to remotely control the HWM 

operation. 

Recent estimates shown in Table 1 (6) suggest that auger and 

highwall mining may account for around 45 million tons of clean 

coal production representing about 4% of the U.S. total production. 

In addition to being a very productive and economic coal 

production method, analysis shows that auger and highwall mining 

appears to be as safe as surface coal mining (6). However, there 

are geotechnical considerations with highwall mining that require 

careful engineering to realize the full economic and safe potential 

of highwall mining. 

Highwall stability is the overriding concern in highwall 

mining. Zipf and Mark (7) discuss the major factors affecting 

highwall stability, namely, geologic structure (i.e., hill-seams and 

mud-seams) and web pillar stability. Unfortunately, little can be 

done to control the location of hill-seams and mud-seams in a 

highwall, and they are difficult to detect reliably. Daily inspection 

of benches above an active highwall mining operation to look for 

cracks and signs of movement is one recommended standard 

operating procedure, easily integrated into the standard pre-shift 

inspection done by a knowledgeable, experienced company official. 

High technology tools such as GroundProbe Inc.’s Slope Stability 

Radar may also prove useful in early detection of movement and 

potential instability in coal mine highwalls (8). Slope Stability 

Radar is presently demonstrating its potential at several open pit 

copper mines in the U.S. (9). Using highwall slope angles in the 70 

to 80 degree range could aid in eliminating much of the stability 

hazard from hill-seams (7). 

This paper concentrates on web pillar stability, a major factor 

controlling highwall stability. The basic equations for computing 

pillar strength and applied stress are presented. By making certain 

assumptions, simple design charts for estimating minimum web and 

barrier pillar widths have been derived, and they are presented 

herein and compared to case history data. These charts apply to 

many routine highwall mining situations. However, exceptions are 

frequently encountered. The case histories are used to illustrate 

many common exceptions, namely 1) the presence of old auger 

holes, 2) highwall mining under back-stacked spoil and finally, 3) 

closely-spaced multiple seam mining. 

Auger and highwall mining are safe and productive surface 

coal mining methods provided that proper geotechnical engineering 

is conducted. Failure to properly engineer web and barrier pillars 

can result in highwall failure, trapped highwall mining equipment, 

and significant economic impact as most of the equipment lies 

under the highwall. These failures result in lost reserves, low 

productivity and potential miner safety issues as personnel work in 

close proximity to the highwall. Recovery of a stuck miner can be 

inherently hazardous since it involves operating under extraordinary conditions. 

ANALYSIS AND DESIGN OF HIGHWALL MINING 

WEB AND BARRIER PILLARS 

Prior work by Zipf (10, 11) derived simple design equations 

for web and barrier pillars. The stability factor for a web pillar is 

Table 1 - Estimated Auger and Highwall Mining Production for 2003 

Machine Approximate number in 

operation 

Productivity 

(raw tons per year) 

Production 

(raw tons) 

Superior Highwall Miners 30 650,000 20,000,000 

ADDCAR Highwall Miners 30 1,000,000 30,000,000 

Augers 150 100,000 15,000,000 

TOTAL (raw tons) 65,000,000 

TOTAL (clean tons) 45,000,000 

Photograph 2. Conveyor car on the launch vehicle of an 

ADDCAR Systems Highwall Miner. 

Photograph 1. “Push beam” being loaded into a SHM. 

SFWP = SI [ 0.64 + 0.54 WWP / H ] (1)

 [SV (WWP + WE) / WWP] 
210

Where: SI = in situ coal strength, 

SV = in situ vertical stress, 

WWP = web pillar width, 

WE = highwall miner cut width, and 

H = mining height. 

In situ coal strength is normally assumed as 900 lb/in2

, unless 

laboratory testing, previous web pillar behavior, coal cleat spacing, 

or site-specific field data suggests a different value to be 

appropriate. Equation 1 is sensitive to in-situ coal strength so care 

must be exercised in assuming a value with site specific testing. 

The mining height is the height of the HMW cut which may be 

greater or less than the seam thickness, dependent upon out-ofseam dilution or if “head” coal is left in the immediate roof to 

improve roof conditions. The type of HWM equipment dictates the 

cut width that varies from 9-ft to 12-ft. In situ vertical stress 

depends on the overlying rock density and overburden depth. 

Vertical stress gradient is typically 1.1 lb/in2

/ft. The high average 

overburden depth can be used to calculate web pillar stress. This is 

the overburden under which 80% of the HWM panel exists or 

alternatively may be calculated using equation 2 where a 

continuous slope is present. 

OBDesign = 0.75 * OBMAX + 0.25 * OBMIN (2) 

Where: OBMAX = maximum overburden depth 

OBMIN = minimum overburden depth. 

For design purposes, the stability factor for web pillars 

typically ranges from 1.30 to 2.00 dependent upon the subsidence 

constraints imposed by current and future surface usage. Based on 

data in HWM ground control plans submitted by coal mine 

operators to MSHA, studies (6) found that the stability factor for 

web pillars ranged from 1.30 to 1.60 in about 30% of the plans and 

exceeded 1.60 in 45% of the plans. The stability factor range 

obtained from the ground control plans is based on statistical 

information provided where the appropriate stability factor is 

unique to each mine site. This survey also found that the width-toheight (W/H) ratio of web pillars exceeded 1.00 in 75% of the cases 

examined. In general, keeping the web pillar W/H ratio above 1.00 

is desirable to maintain web pillar integrity. Narrow web pillars 

increase the demand on surveying accuracy as an orientation 

alignment error of about 4 minutes results in a 1-foot deviation in 

web pillar width in an 800-ft-long cut. For this reason, a minimum 

web width is typically specified irrespective of overburden depth. 

If the number of web pillars in a panel is selected as “N”, then 

the panel width is given by 

WPN = N (WWP + WE) + WE (3) 

A barrier pillar is commonly used to separate adjacent panels 

and prevent ground control problems from cascading along the 

entire length of highwall. Neglecting the stress carried by the web 

pillars (i.e., assuming that they have all failed), the stability factor 

for a barrier pillar is determined as 

SFBP = SI [ 0.64 + 0.54 WBP / H ] (4) 

 [SV (WPN + WBP) / WBP] 

Where: SI = in situ coal strength, 

SV = in situ vertical stress, 

WPN = panel width, 

WBP = barrier pillar width, and 

H = mining height. 

Because the stress carried by web pillars within a panel is 

neglected, the stability factor for barrier pillars can be as low as 

1.00. Studies (6) found that the width of barrier pillars exceeded 

16-ft in more than half the cases examined and more important, the 

W/H ratio for barrier pillars exceeded 3 in 66% of the cases. 

Barrier pillars with a W/H ratio greater than 3 are superior for 

sound geomechanics reasons. 

Based on equation 1, web pillar design charts are developed 

and presented in figures 1 and 2. Figure 1 applies to a 9-ft-wide 

highwall miner cut, while figure 2 applies to a 12-ft-wide cut. In 

figures 1 and 2, options a and b apply to stability factors of 1.3 and 

1.6, respectively. Similarly, based on equation 4, design charts for 

barrier pillars are presented in figure 3. Options a, b and c apply to 

panel widths of 100, 200 and 400 ft, respectively. Note that this 

design chart assumes a barrier pillar stability factor of 1.0 and it 

neglects any load carrying capacity of the web pillars within a 

panel. To use figures 1, 2 or 3, the user begins with the design 

depth on the x-axis, moves up vertically to the applicable mining 

height and then moves left horizontally to the y-axis where the 

suggested web (or barrier) pillar width is read. 

Figure 1A – Suggested web pillar width with stability factor of 1.3, coal strength of 900 lb/in2

0

5

10

15

20

25

100 150 200 250 300 350 400 450 500 550 600

Depth of cover - ft

Web pillar width - ft 

Mining height = 2 ft.

Mining height = 4 ft.

Mining height = 8 ft.

Mining height = 12 ft.
211

Figure 1B – Suggested web pillar width with stability factor of 1.6, coal strength of 900 lb/in2 and 9-ft wide cut.

0

5

10

15

20

25

100 150 200 250 300 350 400 450 500 550 600

Depth of cover - ft

Web pillar width - ft 

Mining height = 2 ft.

Mining height = 4 ft.

Mining height = 8 ft.

Mining height = 12 ft.

Figure 2A – Suggested web pillar width with stability factor of 1.3, coal strength of 900 lb/in2 and 12-ft wide cut.

0

5

10

15

20

25

100 150 200 250 300 350 400 450 500 550 600

Depth of cover - ft 

Web pillar width - ft 

Mining height = 2 ft.

Mining height = 4 ft.

Mining height = 8 ft.

Mining height = 12 ft.

Figure 2B – Suggested web pillar width with stability factor of 1.6, coal strength of 900 lb/in2 and 12-ft wide cut.

0

5

10

15

20

25

100 150 200 250 300 350 400 450 500 550 600

Depth of cover - ft 

Web pillar width - ft 

Mining height = 2 ft.

Mining height = 4 ft.

Mining height = 8 ft.

Mining height = 12 ft.
212

Figure 3A – Suggested barrier pillar width for 100-ft-wide panel assuming coal strength of 900 lb/in2

 and stability factor of 1.0.

0

5

10

15

20

25

30

35

40

45

50

100 150 200 250 300 350 400 450 500 550 600

Depth of cover - ft 

Barrier pillar width - ft 

Mining height = 2 ft.

Mining height = 4 ft.

Mining height = 8 ft.

Mining height = 12 ft.

Figure 3B – Suggested barrier pillar width for 200-ft-wide panel assuming coal strength of 900 lb/in2 and stability factor of 1.0.

0

10

20

30

40

50

60

100 150 200 250 300 350 400 450 500 550 600

Depth of cover - ft 

Barrier pillar width - ft 

Mining height = 2 ft.

Mining height = 4 ft.

Mining height = 8 ft.

Mining height = 12 ft.

Figure 3C – Suggested barrier pillar width for 400-ft-wide panel assuming coal strength of 900 lb/in2 and stability factor of 1.0.

0

10

20

30

40

50

60

70

80

90

100 150 200 250 300 350 400 450 500 550 600

Depth of cover - ft 

Barrier pillar width - ft 

Mining height = 2 ft.

Mining height = 4 ft.

Mining height = 8 ft.

Mining height = 12 ft.
213

HIGHWALL MINING CASE HISTORIES 

The application of web pillar design is examined in the 

discussion of three case histories. The intent of the case histories is 

to illustrate different circumstances under which HWM is 

conducted and the required changes in engineering design to 

accommodate unique mining circumstances. 

Case Study A – Mining Into Old Auger Holes 

In Appalachia, HWM is often conducted in previously augered 

highwalls. Although inexpensive and efficient, auger mining 

sterilizes coal reserves past the typical 50-ft to 150-ft penetration 

depth. The greater 800-ft to 1,200-ft penetration of a highwall 

miner permits recovery of the “island” of coal left in the center of a 

ridge when the perimeter has been augered. 

The concern in HWM through a previously augered highwall 

is that web stability in the initial veneer of augering is a function of 

the auger web widths. The auger webs were either pre-defined in 

the case of multiple head machines or reduced to the minimum 

required to maintain highwall stability during augering. The initial 

challenge is to determine the number of auger webs to be skipped 

between adjacent highwall miner cuts. 

This case history involves re-mining a previously augered 

Stockton seam highwall in Southern West Virginia. The auger 

mining was done using a single head 42-in (3.50-ft) diameter auger 

in the bottom split of the Stockton seam to produce a clean, direct 

ship product. This resulted in leaving a substantial amount of head 

coal given the 7.32-ft to 4.30-ft range of mining height in the 

proposed re-mining area. A portion of the previously auger 

highwall is shown in Photograph 3. 

 

The auger-web-width ranged between 14-in and 22.5-in with 

an average width of 17.6-in. Although the depth of the auger holes 

is unknown, the company engineer said that based upon similar 

mining on the property 150-ft is probably the deepest penetration. 

The stability of the existing auger webs left between the 42-in 

diameter auger holes was examined assuming an average 17.6-in 

width and using in-situ coal strength of 676 lb/in2

 based on prior 

projects in the area. 

Auger web stability was evaluated by considering each web as 

a long, narrow pillar and calculating pillar strength using the well 

accepted Mark/Bieniawski pillar strength formula that defaults to 

equation 1 with the long pillar lengths of HWM. When calculating 

the auger web strength, the 42-in (3.50-ft) auger diameter was used 

for the pillar height although significant “head” coal is present 

above the auger hole. 

However, when calculating auger web stability to determine 

the number of auger webs to skip between adjacent HWM cuts, the 

full seam thickness was used for the mining height. Because the 

HWM cuts the entire seam, it leaves a pillar of the same height as 

the seam. Web pillar stress is calculated using the tributary area 

approach shown in equation 5. Recognizing that a portion of the 

overburden consists of spoil from surface mining of the overlying 

No. 5 Block seam, the overburden density was derated from the 

standard value of 160 to 150 lb/ft3

 or 1.1 to 1.04 lb/in2/ft. The 

actual height of spoil was found to be highly variable and therefore 

difficult to use in the overburden stress calculation. The reduced 

overburden density is a compromise between using the standard 

rock density and the spoil and intact overburden heights. 

 Sp = 1.04(H)(w + B)(l + B) (5) 

 (w)(l) 

Where: Sp = pillar stress (lb/in2

), 

H = overburden depth (ft), 

w = pillar width (ft), 

B = entry or crosscut width (ft), and 

L = pillar length (ft). 

The auger hole depth was not accurately recorded at the time 

of mining. Consequently, an auger web length of 150-ft was used 

since this is likely the greatest penetration achieved with a single 

head auger. The previously augered highwall was divided into 

individual sub-areas based upon seam thickness and overburden 

depth prior to conducting the auger web stability calculations. The 

auger web stability calculations were initially performed by 

considering only the auger web and therefore reported in terms of 

lb/in2

. When calculating the number of web pillars to skip between 

adjacent HWM cuts, it was easier to determine the tons of 

overburden load placed on both the HWM cut and the augered area. 

Similarly, the bearing capacity of the aggregate of the auger webs 

was reported in tons. 

 The planned re-mining was done using a HWM machine that 

makes an 11.50-ft-wide cut into the highwall. The 11.50-ft-wide 

cut encompassed two 42-in-diameter auger holes and a 17.6-in 

average auger web, plus it exposed the next auger hole leaving a 

void space of 13.43-feet. Therefore, all HWM calculations were 

done assuming a 13.43-ft-wide void and not the standard 11.50-ftwide HWM cut. 

The web pillar stability factors for each sub-area were 

presented in table 2, which documents the stability factors for the 

existing auger webs from a nominal 20-ft of overburden to the 

maximum overburden depth present within the assumed 150-ftpenetration depth of the auger hole. Table 3 shows the number of 

auger webs that should be skipped between adjacent HWM cuts. 

The boldface and shaded areas of tables 2 and 3 are those 

combinations of web width and overburden thickness that are not 

sufficiently stable to insure the web stability. 

Photograph 3. A Previously Augered Stockton Seam Highwall.
214

 By reviewing Table 3 for each sub-area, it is readily apparent 

that the ability to re-mine the existing auger holes is limited by the 

stability of the existing auger web pillars. A threshold SF of 1.30 is 

used to separate mineable from marginal areas. It should be noted 

that a stability factor of 1.30 is low and is for short-term stability 

only. It is clear from Table 3 that the ability to skip auger webs to 

maintain highwall stability rapidly diminishes at greater than 100-ft 

of overburden. A skip of one HMW cut is recommended between 

each group of 10 HWM cuts to provide a barrier pillar and prevent 

a squeeze in previously mined HWM cuts and auger webs from 

cascading onto the active group of HWM cuts. 

 The mining was successfully done by a contractor who did 

initially adhere to the recommended number of auger web skips 

between adjacent HWM cuts. As the job progressed and the 

contractor gained experience with the behavior and response of the 

auger webs to HWM, the number of skipped auger webs was 

selected based upon a combination of engineering calculations and 

site conditions. 

Case Study B - Mining Under Back-Stacked Spoil 

 A highwall web pillar collapse occurred in the 11th cut of a 

peninsula-shaped remnant left after Pittsburgh No. 8 seam was 

contour surface mined in the 1970’s. The reclaimed surface mine 

was re-opened for HWM. The spoil was removed and a second 

contour cut was taken to provide a clean, stable highwall under 

which to face up the Southwest side of the peninsula and operate a 

HWM. HWM web pillar widths were designed using the 

anticipated overburden range, mining height, and a 900-lb/in2

 insitu coal strength. A 75-ft to 80-ft high spoil pile from the second 

cut was placed on top of the peninsula, directly above where the 

highwall miner was to begin cutting. Starting from the Northwest 

side, eleven highwall miner cuts were driven into the peninsula 

from the new face-up. 

 In cuts 1 and 2, the highwall miner operator attempted to stay 

below the Pittsburgh No. 8 Rider seam that is approximately 1-ft to 

1.5-ft above the Pittsburgh No. 8 seam main bench. The 

interburden is very weak and the immediate roof strata fell in upon 

cutting. In subsequent cuts (3 through 7 and 9 though 10), the 

interburden was taken to the height of the Pittsburgh No. 8 Rider 

Table 2. Auger Web Width Stability. 

Stockton Seam - Pillar Safety Factors - Existing Single Head 42 Inch Diameter Auger Holes 

Bieniawski and Mark/Bieniawski Pillar Strength Formulae 

Auger 

hole 

diameter 

B (feet) 

Web 

length l 

(feet) 

Web 

width w 

(feet) 

Coal 

seam 

height h 

(feet) 

Web 

pillar 

height h 

(feet) 

Overburden 

depth (feet)

Overburden

Stress 

(lb/in2

) 

Bieniawski 

pillar strength

formula 

(lb/in2

) 

Bieniawski 

safety 

factor 

Mark/ 

Bieniawski 

strength 

formula (lb/in2

)

Mark/ Bieniawski

safety factor 

Auger Hole Stability - Areas 50-4, 27-3, & 27-2 - 17.58 Inch Web Thickness Between 42 Inch Diameter Auger Holes 

3.50 150 1.47 5.00 3.50 20 70 535 7.58 585 8.30 

3.50 150 1.47 5.00 0.50 40 141 535 3.79 585 4.15 

3.50 150 1.47 5.00 3.50 60 211 535 2.53 585 2.77 

3.50 150 1.47 5.00 3.50 80 282 535 1.90 585 2.07 

3.50 150 1.47 5.00 3.50 100 352 535 1.52 585 1.66 

3.50 150 1.47 5.00 3.50 120 423 535 1.26 585 1.38 

3.50 150 1.47 5.00 3.50 140 493 535 1.08 585 1.19 

Marginal Safety Factor 

Table 3. HWM Web Pillar Stability and the Number of Auger Webs Skipped Between HWM Cuts. 

Stockton Seam - Pillar Safety Factors - Number of Skipped Auger Webs Between Highwall Miner Cuts 

Bieniawski and Mark/Bieniawski Pillar Strength Formulae 

Highwall 

miner 

cut b' 

(feet) 

Auger 

hole 

diameter 

b (feet) 

Highwall 

miner cut 

length l 

(feet) 

Web 

width w 

(feet) 

Coal 

seam 

height 

h (feet) 

Number 

of web 

pillars 

between 

HWM

cuts 

Overburden

depth (feet)

Overburden

load (tons)

Bieniawski 

pillar 

strength 

formula 

(tons) 

Bieniawski 

safety 

factor 

Mark/ 

Bieniawski 

strength 

formula 

(tons) 

Mark/ 

Bieniawski 

safety factor

Area 50-4, 27-3, & 27-2 - Number of Auger Webs Between Each HWM Cut 

13.43 3.50 150 1.47 5.00 4 20 7,491 31,902 4.26 34,137 4.56 

13.43 3.50 150 1.47 5.00 4 40 14,981 31,902 2.13 34,137 2.28 

13.43 3.50 150 1.47 5.00 5 60 25,823 39,877 1.54 42,671 1.65 

13.43 3.50 150 1.47 5.00 8 80 47,837 63,804 1.33 68,274 1.43 

13.43 3.50 150 1.47 5.00 15 100 98,898 119,632 1.21 128,014 1.29 

13.43 3.50 150 1.47 5.00 15 120 118,677 119,632 1.01 128,014 1.08 

13.43 3.50 150 1.47 5.00 15 140 138,457 119,632 0.86 128,014 0.92 

Marginal Safety Factor 
215

seam. Intermittent falls occurred in the Pittsburgh No. 8 Rider 

seam exposing a weak overlying gray shale unit. Once the gray 

shale is exposed to the atmosphere, the shale degrades and roof 

falls are propagated upward as an arch and terminated against a 

competent brown sandstone bed. 

Full penetration of 800 ft was achieved in one cut because roof 

falls of various magnitudes occurred in each cut. The eleven cuts 

ranged in depth between 362 and 810 ft. Difficult roof conditions 

were routinely encountered in each cut within a zone from 400-ft to 

600-ft deep. Cut 8 was skipped because of a highwall slide. Cut 

10 reached a depth of 810-ft when a roof fall terminated further 

advance. The highwall miner head had advanced to a depth of 710-

ft in cut 11 when a web pillar collapse and roof fall trapped the 

machine. 

The layout of the eleven cuts is shown above in figure 4. The 

arched roof falls are clearly seen in Photograph 4 that shows the 

condition of highwall miner cuts 9, 10, and 11. The tail end of the 

push beams can be seen in cut 11 on the right side of the 

photograph. 

 

The cause of the web pillar collapse was 

 

 back-stacked spoil which increased the web pillar 

loading not being incorporated into the web pillar design 

and 

 an increase in mining height attributable to the roof falls 

that reduced the width/height ratio of the web pillar and 

ultimately the web pillar strength 

The following recommendations were offered to reduce the 

likelihood of web pillar failures in subsequent mining. 

 Mining from a fresh second contour strip cut and scaled 

face-up with a minimum height of 60 ft, 

 Leaving sufficient “head” coal from the main bench of 

the Pittsburgh No. 8 seam to provide support for the 

overlying interburden, Pittsburgh No. 8 Rider, and 

immediate roof strata, 

 Not placing spoil or unconsolidated material on the 

highwall above areas where mine personnel work or 

travel, 

 Avoid mining in areas of low overburden (<60 ft) 

thickness, 

 Orienting the highwall miner cuts to start from highest 

overburden and mine to the lowest overburden, and 

 Maintaining a standard 6-ft-wide web in the Pittsburgh 

No. 8 seam, unless highwall mining is or will be present 

in the Meigs Creek No. 9 seam, at which point the web 

width should be increased to 8.75 ft. 

 

Case Study C – Closely Spaced Multiple Bench Mining 

This case history concerns a contract HWM operator mining 

the Stockton coal seam from surface contour pits developed by the 

parent mining company. In the active mining area, the Stockton 

seam occurs in two splits, ranging between 42-in to 45-in thick, 

separated by an in-seam parting. The in-seam parting, as seen in 

the surface mine pits, is of variable thickness and consists of a 

competent sandy fireclay (Ferm No. 327) and weak shaley coal. 

Samples of the upper split of the Stockton seam and the sandy 

fireclay parting were recovered as part of the site inspection. The 

uniaxial compressive strength of the coal and the tensile strength of 

the in-seam parting are the basis for web pillar and in-seam parting 

stability factor calculations. 

In-seam parting stability was analyzed using the simply 

supported beam equation (6) to determine the required in-seam 

parting thickness so that the in-seam parting will remain stable 

while the lower split is safely extracted. The simply supported 

beam equation was selected because it assumes that the beam ends 

can flex and is appropriate for lower (<300 ft) overburden depths 

Peng, (12). No strength was assigned to the shaley coal unit 

because of the variability of this stratum. The in-seam parting 

strength and stability factor calculations are based on the sandy 

fireclay. The tensile strength and density of the sandy fireclay is 

590 lb/in2

 and 169.50 lb/ft3. 

σt = 3γL2

 (6) 

 4t 

Where: σt = tensile strength (lb/in2

), 

γ = density (lb/ft3

), 

L = opening width (ft), and 

t = beam thickness of individual stratum (ft). 

Figure 4. Mine Layout Where HWM was Trapped. 

Photograph 4. Roof Falls and Web Collapse 
216

In the upper section of table 4, the stability factor for a 

standard 10.40-foot wide opening is shown along with the 

maximum allowable highwall miner opening for stability factors of 

1.00 (marginal stability), 4.00 (short-term stability), and 8.00 (longterm stability) given the actual in-seam parting thickness, and the 

tensile strength and density of the parting material. 

The lower portion of the table is a sensitivity analysis where 

the same analysis is conducted for a range of in-seam parting. 

Obert and Duvall (13) define the stability factor criterion of 4.00 

and 8.00 “for members in tension such as bedded roof.” Those 

combinations that do not meet this criterion are marginal and are 

shown as shaded. 

It is clear, based upon Table 4, that the combination of inseam parting thickness and the 10.4-ft HWM cut width present in 

the previous pit do not satisfy the long-term stability criterion and 

in one instance does not meet the short-term requirements. A 

minimum sandy fireclay parting thickness of 1.50 ft is necessary to 

provide long-term stability when mining both splits of the Stockton 

seam.

Since the parting lithology is subject to change across the 

property, the minimum thickness applies to competent shales (Ferm 

No. 124), sandy shales, or sandstones. Caution was strongly 

recommended if the parting consists of weak strata, for example, 

fireclay (Ferm No. 127), laminated sandstone, shale with coal 

stringers, or fractured strata. 

The obvious question from Table 4 is why did the failure 

occur at HWM cut number 16 when the in-seam parting stability 

factor was greatest and appears to have exceeded the short-term 

stability threshold. In conversation with the operator during the site 

inspection, he believed that the HWM cut orientation deviated so 

the openings in the upper and lower split were not columnized. 

Stability in multiple seam or multiple split mining is dependent 

upon the transfer of overburden stress through the upper web pillars 

to the lower web pillars. If a HWM cut is not perfectly aligned 

with the overlying cut, the overburden stress is transferred onto the 

in-seam parting. In this situation, failure is likely to occur as the 

web punches through the in-seam parting. The stability factors in 

table 4 reflect only the ability of the in-seam parting to remain selfsupporting, not its ability to resist vertical loads from an overlying 

web. When there is any chance that the webs will not be perfectly 

columnized, a tensile failure analysis similar to one described by 

Zipf (11) is recommended. 

CONCLUSIONS 

Highwall mining is an efficient and economic means of 

surface mining. In contrast to auger mining, HWM requires a 

thorough knowledge of the strength and physical properties of the 

immediate roof, coal, and immediate floor strata. Engineering 

design and surveying the orientation of each HWM cut is necessary 

to avoid ground control problems. Ground control is critical in 

HWM because a significant portion of the $5MM to $6MM HWM 

machine lies beneath the hillside with virtually no easy access to 

resolve a roof fall or pillar squeeze. 

The analytical design equations provided in this paper present 

the base from which web pillar design should begin. However, as 

illustrated by the case histories, site conditions and characteristics 

of the coal seam and immediate roof strata frequently require 

deviation from the web pillar width calculated using the 

Mark/Bieniawski equation. 

Table 4. Simply Supported Beam Stability factors for in-seam parting between upper and lower splits of the Stockton seam 

Simply support beam calculations HWM hole 

number 

Parting 

thickness 

(feet) 

Tensile 

strength 

(lb/in2

) 

Density 

(lb/ft3

) 

Simply supported 

beam safety factor 

10.4 ft. Span 

Maximum roof span

for a S.F. = 1.00 

Maximum roof span 

for a S.F. = 4.00 

Maximum roof span

for a S.F. = 8.00 

Safety Factors for Actual Sandy Fireclay In-Seam Parting Thickness in the Previous Pit 

16 1.35 520 169.50 7.35 28.19 14.09 9.97 

17 1.06 520 169.50 5.77 24.98 12.49 8.83 

18 1.30 520 169.50 7.08 27.66 13.83 9.78 

19 1.20 520 169.50 6.53 26.58 13.29 9.40 

20 1.20 520 169.50 6.53 26.58 13.29 9.40 

21 1.09 520 169.50 5.93 25.33 12.67 8.96 

22 0.84 520 169.50 4.57 22.24 11.12 7.86 

23 0.90 520 169.50 4.90 23.02 11.51 8.14 

24 0.60 520 169.50 3.27 18.79 9.40 6.64 

Safety Factors for a Range of In-Seam Parting Thickness 

 3.00 520 169.50 16.33 42.02 21.01 14.86 

 2.50 520 169.50 13.61 38.36 19.18 13.56 

 2.00 520 169.50 10.88 34.31 17.16 12.13 

 1.75 520 169.50 9.52 32.10 16.05 11.35 

 1.50 520 169.50 8.16 29.71 14.86 10.51 

 1.25 520 169.50 6.80 27.13 13.56 9.59 

 1.00 520 169.50 5.44 24.26 12.13 8.58 

 0.75 520 169.50 4.08 21.01 10.51 7.43 

 0.50 520 169.50 2.72 17.16 8.58 6.07 

 0.25 520 169.50 1.36 12.13 6.07 4.29 
217

REFERENCES 

1. Volkwein, J.C., Mucho, T.P. and Bhatt, S.K. Safety, 

Technological, and Productivity Potentials of Highwall Mining. 

U.S. Dept. of Labor, MSHA, Holmes Safety Association 

Bulletin, January 1995, pp. 4-11. 

2. Arrowsmith, D. Highwall Evolution. World Coal, Vol. 12, no. 

9, Sept. 2003, pp. 15-18. 

3. Fiscor, S. Contour Mining Puts Highwall System to the Test. 

Coal Age, July 2002, pp. 20-22. 

4. Superior Highwall Miners, L.P., Beckley, West Virginia, 

http://www.shm.net, 2004. 

5. Mining Technologies, Inc., Ashland, Kentucky, 2004, 

http://www.addcarsystem.com. 

6. Zipf, R.K. and Bhatt, S.K. Analysis of Practical Ground 

Control Issues in Highwall Mining. In Proceedings, 23rd 

Conference on Ground Control in Mining, West Virginia 

University, Aug. 3-5, 2004, pp. 210-219. 

7. Zipf, R.K. and Mark, C. Ground Control for Highwall Mining 

in the U.S.,” International Journal of Surface Mining, 

Reclamation and Environment, accepted for publication. 

8. GroundProbe Pty. Ltd., Brisbane, Queensland, Australia, 

http://www.groundprobe.com, 2004. 

9. Seppala, V. Integration of Technology in Slope Management 

Programs. SME Annual Meeting, preprint 05-34, 2005, 10 pp. 

10. Zipf, R.K. Catastrophic Collapse of Highwall Web Pillars and 

Preventative Design Measures. In Proceedings, 18th 

Conference on Ground Control in Mining, West Virginia 

University, Aug. 3-5, 1999, pp. 18-28. 

11. Zipf, R.K. Ground Control Design for Highwall Mining. SME 

Annual Meeting, pre-print number 05-82, 2005, 9 pp. 

12. Peng, S.S. Coal Mine Ground Control. John Wiley & Sons, 

New York, New York, 1978, pp. 124-127. 

13. Obert, L. and Duvall, W.I. Rock Mechanics and the Design of 

Structures in Rock. John Wiley & Sons, New York, 1967, 

pp. 490.