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By Yang Li

For the study of overburden movement in longwall mining of shallow coal seams, most scholars merely consider the thickness of bedrock and top soils. In this paper, the overburden movement in longwall mining of shallow coal seams with thin top soils and key strata group was studied. By using numerical simulation, the development of fractures in the overburden was analyzed for different ratios of panel width to depth (W/D) and depth to mining height. The results show that the overburden movement changed with the variation of mining height, panel width and depth under the specific geological conditions. It also explain how different overburden disturbance structures, such as ?three zones? (caved, fractured, and continuous deformation zones), ?two zones?(caved and fractured zones) or ?steps-like collapse? are formed under different mining conditions. The results broaden the vision for research on overburden movement in longwall mining of shallow coal seams.

Jan 1, 2014

By Brian Daniels, James Farroni

"Draw rock (small rock that forms/falls when bolting the roof and/or rib) is the cause of numerous injuries annually. These injuries result in lost work hours and low morale. The report below explains how ""Mine Safe Draw Rock Shields"" (Pans) are used in an underground operation. This report will chronicle the application of these pans in a Southern West Virginia coal mine along with testing data that measures their effectiveness. Since the introduction of these pans in 2002 the amount of reportable accidents relating to draw rock has decreased by 80%. The report below reinforces that perimeter contact is critical to the performance of these pans. In addition the strength of these products at the perimeter will determine their effectiveness. The results of the testing data outlined in this paper, along with the physical records at the study mine, confirm that the use of pans has significantly reduced draw rock falls at this facility.BACKGROUNDSupplemental support is defined as ground control that is used in addition to required support in an underground mine. Some examples of supplemental support are:• Pans• Mesh/Geogrid• Mats• Half Headers• Boards• ChannelA typical roof condition that warrants the use of pans is when loose small rock is dislodged while bolting. If the top is very scaly (loose, thin rock) additional supplemental support can be used in conjunction with pans such as mesh or geogrid (See Figure 1). Pans will provide protection where used and approximately 12 in (250 mm) out from perimeter. This is achieved by contact with a surface which will put scale in tension up to 12 in (250 mm) from perimeter of pan.INTRODUCTIONThe ease of applying a roof support system is critical to effectively operating a deep mine. Figure 2 shows how a pan is installed in conjunction with a roof bolt and bearing plate. The Mine Safety and Health Administration (MSHA) requires that supplemental support cannot decrease the performance of required support. The goal of using all supplemental support is to provide additional safety for the miner which equates to increased productivity.The ability of skin control products to contact a surface is the single most critical characteristic (Figure 3 and Figure 4). If the material does not contact the surface it cannot be effective. Therefore, the flexibility, design and proper installation of the products are critical in determining their effectiveness. Figure 5 shows the proper installation of the roof pan.To reiterate, all skin control must be applied in conjunction with a roof bolt and ASTM approved bearing plate as shown in Figure 4."

Jan 1, 2010

By B. (Mamas) Das

A project to evaluate the suitability of point Load strength tests for testing and classifying coal measure rocks for coal mining application was initiated by Canada Centre for Mineral and Energy Technology (CANMET). A summary of this study with emphasis on the important findings is presented in this paper. The instrument as used in this study is a Robertson Research Rock Strength Index Log. Because of the variation in the correlation ratio between uniaxial compressive strength and point Load strength index for hard rocks and soft rocks it was decided to establish this corre¬lation initially on cement concrete specimens specially prepared for this purpose following ASTM specifications. These specimens are expected to represent soft rocks and at the same time would be of uniform quality. Once a general trend of this correlation was established tests were performed on coal and rock samples from the sub-bituminous and bituminous coalfields in Canada. The most important results may be stated as follows: (i) Generally, the common value of the ratio between the uniaxial compressive strength and the point load strength index ranges between 8 to L4. (ii) The Robertson Research Rock Strength Index Log proves to be quite suitable for testing sandstone, siltstone and some coals from the bituminous coalfields but problems have been experienced when testing weathered samples or samples from the sub-bituminous coalfields. (iii) In some cases tests were successful if Larger samples from weathered zones or from sub-bituminous coalfields were used. However, a magnified load gauge for 0-3KN or 0-5KN is expected to be the best solution for soft and weathered rocks in most of the cases. Some mudstones and some rocks from weathered zone do not appear to be suitable for these tests. This aspect of the study is still under investigation. (iv) Deterioration of rock properties due to weathering has been determined through tests on samples collected from weathered and unweathered samples.

Jan 1, 1984

By Francis S. Kendorski

Underground stone mines tier construction aggregates and industrial minerals are being planned in increasing numbers throughout the United States. At the middle of the last century (1950s) underground stone mines were much more common, and characteristically had no thorough site characterization other than proving reserves and trial-and-error planning. Ground control employed was minimal. Consequently, there was a tendency to conduct minimal pre-mining studies. However, the deposits that lend themselves to little site characterization and minimal ground control are few today, with the best being mined out. Failures of a number of underground stone or similar mines from that era are more common than realized, Underground stone mines almost always use some combinations of square or rectangular pillars, sized to support the overburden load, and spaced to allow the use of larger, more-efficient equipment. Heights are controlled by the available stone quality, interbeds of waste, or technological or safety considerations. Prior to committing to a specific mine plan with subsequent significant capital investment, and prior to having available the wealth of knowledge and experience gained once mining underground, a safe and workable mine layout must he achieved. Early attention to the several different rock units involved (roof, mine level. floor) and their strength as a rock mass with the inherent discontinuities, will allow first approximations of acceptable mine geometries and sequencing.

Jan 1, 2000

Effective strata control, utilising fully encapsulated roof bolts, is dependent on the installed quality of the reinforcement elements. One mechanism by which fully encapsulated roof bolts may become less than fully effective is by gloving (glove fingering) and un-mixing of the resin. Following some small roof failures containing gloved bolts, investigations have been undertaken in some New Zealand coal mines to determine the extent of, and mechanisms involved in, the gloving of fully encapsulated roof bolts for a range of roof types and installation methods. Gloving, in this context, refers to the resin cartridge partially or completely enveloping a length of the bolt, typically with unmixed resin filler and catalyst/hardener remaining within the cartridge. Results have shown that gloving and un-mixing is a systematic and widespread phenomena, occurring across a range of resin and/or bolt manufacturers, and across a variety of roof types. Gloving was found in bolts installed using: hand held pneumatic and continuous miner mounted hydraulic methods, under run of mine (ROM) conditions by mine workers, and under controlled "best practice" conditions. The roofs bolts investigated are all 1800mm (6ft) long, and have shown that, on average, 485mm (-19") of the bolt length is gloved and/or unmixed. This equates to a 27 percent reduction in effective bolt length. The effected length of bolt is sufficient to impact on the load transfer/reinforcement capabilities of the bolt system, and hence impacts on the design assumptions and subsequent stability assessments. Furthermore, there is a direct relationship between the mechanism for the gloving of bolts and an increased occurrence of resin loss and reduction in encapsulations length in soft (less than 10 MPa) and/or highly fractured rock. Investigations have concentrated on developing an understanding of the mechanisms involved. This has been undertaken using both field installations and recovered bolts, as well as test bench trials. The mechanisms involved have been confirmed as being: the development of a pressure front as the bolt encounters the resin and is spun up the hole, which, in turn, leads to over-pressurisation and radial expansion of the resin cartridge, resulting in an increase in diameter of the plastic resin cartridge, which allows the bolt to be spun up into the cartridge without making sufficient contact to shred the cartridge or the hardener tube, resulting in an unmixed portion of resin.

Jan 1, 2003

By David K. Ingram

The objective of this U.S. Bureau of Mines (USBM) investigation was to characterize overburden response due to longwall mining. Subsurface strata and surface deformations were monitored at two sites in southeastern Ohio. Site A included two adjacent longwall panels. Site B consisted of one longwall panel. Both sites had similar mine geometries and overburden thicknesses. The major difference between the two sites was the rock strata within the overburden. Multipoint borehole extensometers (MPBX) were employed to measure subsurface displacements. Survey monuments were installed above the centerline and profile of each panel to measure surface deformations. Subsidence of the surface was observed before and after movement was detected in the MPBX units. Initial and final displacement of the MPBX units occurred at the same time at both sites. No deformation was detected outside the panel area. The development of overburden deformation was different over each panel. The final subsurface failure characteristics also varied between panels.

Jan 1, 1994

By R. Karl Zipf

Highwall mining is an important coal mining method. Upwards of 60 highwall miners are presently in operation, and they may account for approximately 4% of total U.S. coal production. A review of the Mine Safety and Health Administration (MSHA) accident data over the 20 year period from 1983 to 2002 identified 9 fatalities attributable to auger and highwall mining of which inadequate ground control accounted for 1/3. In the past 5 years, 1 fatality occurred in highwall mining. Estimates of the manpower requirements in highwall mining suggest that its fatality rate is essentially the same as for surface coal mining, thus highwall mining appears to be a very safe modern mining method. Highwall stability is the major ground control related safety concern, and operators are required to develop and follow an appropriate highwall mining ground control plan. The plans usually specify the following geotechnical parameters: hole width, maximum hole depth, maximum overburden depth, seam thickness, web pillar width, barrier pillar width and number of holes between barriers. Calculated web pillar stability factor exceeded 1.3 for most designs evaluated. This study examined records from 5,289 highwail miner holes with a total completed footage of about 2,560,000 feet to understand the reasons for early pull out. Average lost footage is typically about 20% of planned footage. Only 35% of the holes reached planned depth, and 20% were short due to rockfalls. Water and adverse geology accounted for 15% of the losses. Mechanical/electrical problems, guidance, and slope stability problems accounted for the remaining 30%. Web pillar stability factor for these holes also exceeded 1.3 in 95% of cases. Trapped highwall miners present a safety hazard to workers during their recovery. Best practices to avoid trapped highwall miners include: avoid mining in stream valleys, avoid mining near outside corners, careful alignment of each hole and the use of an onboard guidance system, Several issues in highwall mining ground control require further investigation including highwall mining through old auger workings, highwall mining near old underground mines, multiple-seam and multiple lift highwall mining and finally the size and frequency of barrier pillars.

Jan 1, 2004

By Zhanjing Yu

In this paper, an equation is developed to characterize the relationship between the UBC and plate size using a boundary integral approach. The equation was validated by the results from mo& than seventy (70) plate loading tests of different sizes conducted in various underground coal mines throughout the Illinois Coal Basin. The equation indicates that the UBC decreases as the plate size increases. The fundamental solution for this boundary integral problems is the Boussinesq equation. The load exerted by the plate on the floor surface is assumed to be uniformly distributed. Both normal and tangential loads are considered. It is also assumed that the floor fails in shear in a plate loading test. The equation may provide the mine operators with a useful tool to obtain from a plate load test of any plate size the UBC value required by the regulations.

Jan 1, 1993

By S. M. Hsiung

A model (support selection model) for the selection of suitable types of powered supports under various geological conditions is proposed. This model is developed by evaluating the support, performance under different roof conditions at longwall faces in terms of five controlling parameters: (1) roof Falls, (2) front abutment pressure, (3) Face convergences, (4) gob materials entering into working area, and (5) roof caving line. This model suggests that more than one type of powered support. may be used under the same roof condition and the best selection should he based on the magnitude of suitability index of each support- type. The development and application of suitability index arcs discussed. Fourteen panels in seven coal seams are discussed in the paper to illustrate the applicability and effectiveness of this model. A roof classification system at longwall faces developed specifically for the construction of the selection model is also discussed. A simple example fur determining roof classification index is given.

Jan 1, 1988

By Russell C. Frith

Risk-based roadway roof support design is now a critical part of the Australian Coal Industry. Safe and efficient mining demands that roof support be tailored to the prevailing geotechnical conditions and coal mining legislation in New South Wales, Queensland, and Tasmania is clear in requiring that formalized roof support design be undertaken. The assumption of dead-load suspension of an otherwise unstable roof has been used in roadway roof support design for many years. However longevity does not necessarily translate into either a best or even reasonable practice in the current industry, particularly when pro-active strata management systems are now routinely used and reinforcing support design methods are available. The paper discusses why the assumption of dead-load suspension is fundamentally incorrect in almost all instances when pro-active or reinforcing roof support is applied and how it can easily result in misleading and potentially under-designed roof support systems under various circumstances. The commonly used design methodology of balancing the installed axial capacity of long tendons with the assumed ?weight? of a future roof fall is summarized, and several fundamental flaws are identified. The critical area of selecting the design Factor of Safety is also discussed in detail. Accepting that under certain circumstances the roof of a roadway will require it to be ?suspended?, suggestions are given for how a robust support system can be developed, designed and applied.

Jan 1, 2011

By Xicai Gao

With the intensification and expansion of deep coal mining, the surrounding rock stress environment becomes further deteriorated because of the complicated geological environment and powerful mining effect, so the major deformation of roadway appeared frequently. The proportions of floor heave are increasing and , the most serious problem is the floor heave of preparatory workings. The coal seam of Binchang coal area in Shaanxi province is at a depth of 621.6 ~ 817.0 m, the average thickness of the coal seam is 25 m, and the overburden and coal measure strata are the main regional aquifers and are highly permeable. The properties of water-weakening are obvious in the floor strata. The main preparatory workings or chambers were driven in the coal seam. The severe floor heave of the water chamber appeared during early mining and the amount was 180 ~ 910mm, and seriously affected the normal installation and use of the water chamber. Theoretical study and numerical simulation as well as the in-situ measurements were carried out to investigate the deformation and failure characteristic of the surrounding rock with non-support in the floor and the main reasons which caused the floor heave were determined. A new combined support technology was proposed which consists of overcutting, grouting, backfilling and invertedarch and bolt-grouting. The field monitoring showed that the cumulative convergence of the water chamber was 20mm, the maximum convergence speed is 1.2 ~ 2mm/d, the accumulative convergence of two ribs was 36mm and the maximum convergence speed is 2mm/d. The combined support technology will enhance the support intensity between the roof and floor, the two sides of the entry and improve the loading capacity of the floor effectively and the floor heave can be controlled effectively.

Jan 1, 2013

By Tom Blevins, Roland Walker, Jinsheng Chen

"The Dywidag-Systems International DSI Tandem Anchor bolt, with an improved expansion shell-washer design, can substantially reduce potential entry roof falls caused by twisted or deformed expansion shells of the traditional mechanical anchor bolts during installation. In addition, the improved anchorage system can also ensure that a consistent and high installed load is achieved, which is critical for effective beam building under thinly laminated or layered roof conditions. This paper discusses the genesis and anchorage performance of the Tandem Anchor bolt, and analyzes the tension/torque relationship of different mechanical shell anchors using data obtained from both lab and field tests. It also discusses roof bolting mechanisms, roof bolting systems, and factors influencing effective beam building. Finally, numerical modeling was used to analyze the effect of bolt pretension magnitude on entry roof stability.INTRODUCTIONThroughout the history of underground mining, keeping the roof from uncontrollable collapse has been the greatest concern of the miner. Tremendous strides have been made in better understanding, predicting, and supporting the mine roof. This conference has spent the last twenty-eight years sharing information and distributing it to the mining community. A paraphrased statement from former Secretary of Defense Donald Rumsfeld adequately describes the challenges to the mining engineer as it pertains to roof control: it is a study of known knowns, known unknowns, and unknown unknowns. It is the hope of the authors to address some of these unknowns.The paper discusses the development of bolt anchorage systems from dry shells to shells with resin-assisted anchorage. It looks at some of the issues faced with the anchorage shells being used when introducing the insertion pressures caused by resin cartridges. The rationale behind the shell design changes was to compensate for this resin induced stresses.An improved expansion shell assembly (a slotted expansion shell base, a tapered cramming plug, and a novel support nut with tapered top) for mine roof bolts makes up the Tandem Anchor (U.S. Pat. No. 6742966); this minimizes the possibility of twisting seen with traditional mechanical shells and thus prevents springy bolts and shell leaves from breaking. A fully grouted Tandem Anchor bolting system has the advantages of both the pure mechanical point-anchor and the non-pretention fully grouted bolting systems. Once installed, the installed load or pretension will be locked in and never bleed off. Compared with the fully-grouted (2-speed resin) tension rebar bolting system, the fully-grouted Tandem Anchor bolting system can (a) reduce bolting cycle time (about 15 seconds per bolt), boosting coal production and productivity; (b) provide a greater installed load or pretention; and (c) achieve a consistent tension/torque ratio, which gives better bolting quality control. Therefore, a fully grouted Tandem Anchor bolting system is the most effective means to reinforce a weak rock mass intrinsically by the theory of beam building. For thinly laminated or layered roof conditions, a higher installed load or pretention (50% or more of the bolt's yield strength) can make a roof bolting system much stiffer, which results in less roof movement. Numerical modeling results show that the magnitude of the roof's tensile stress and relaxation zone can be substantially reduced under a higher installed load or pretention."

Jan 1, 2010

By John Latilla

Following its introduction on underground collieries, the mandatory Code of Practice to Combat Rockfall Accidents (CoP) was applied to all opencast operations of lngwe Collieries (the South African division of BHP Billiton Energy Coal). The CoP includes, but is not limited to, requirements to assess hazards, manage them and monitor compliance. Information on methods to measure highwall conditions and compliance with control measures on opencast strip mines, as opposed to open pit mines, was found to be anything but freely available. However, on the Ingwe underground mines, a system of rating physical conditions in each section and how well production personnel were managing the situation, has been used with success for some time. It was decided that a similar approach would be taken, and, as with the underground rating system, the resultant rating forms should not be aimed at rock engineering specialists only but should be able to be used by production personnel as well. This paper discusses the highwall risk rating method as evolved over a couple of years and highlights that it is a flexible tool that can easily be changed to suit an individual mine's requirements. It also shows how operational aspects are being measured to determine how well the identified hazards are being managed. Allied to the method is a relatively easy, and certainly "low capex," mapping technique which allows between 500 and 1000m of highwall to be covered in a day. Features generally recorded include lithology, bench height, work still to be completed, coal losses or "toes", presence of water, dolerite intrusions, existing and potential failures, prominent joints, declared special and cautionary areas, cracks on upper benches and condition of spoil piles. Highwall mapping and photography as well as basic geological data collection, enable highwall risk and operational ratings to be determined. The same rating sheets may also be completed by production personnel, but without detailed mapping.

Jan 1, 2002

By James M. Tennant

In the mid 1970's, American Electric Power significantly expanded its internal coal production. This involved opening several large mines in seams that had never been mined on a large commercial scale before. One of the mines opened was the Martinka No. 1 Mine which is located approximately five miles southeast of Fairmont. The mine was originally designed for room and pillar mining with continuous miners but was later modified to permit the utilization of longwalls and continuous miners for entry development. The coal seam to be mined was the Lower Kittanning which averages fifty-six inches. In order to successfully mine this seam, it required that the mine design recognize the effect caused by the dip of the seam which averages four percent to the Northwest with a maximum of sixteen percent for short distances and a relatively low cover. In fact, seventy-four percent of the reserves have cover of less than four hundred feet. Also, the core hole data indicated that the immediate roof varied from zero to eighteen feet of shale with two to forty-five feet of sandstone overlying the shale.

Jan 1, 1982

By Swapan Bhattacharya

Section 516 of Public Law (PL) 95-87, commonly known as the Surface Mining Control and Reclamation Act of 1977 (SMCRA), requires that underground coal mine operators must 'adopt measures consistent with known technology in order to prevent subsidence causing material damage to the ex- tent technologically and economically feasible, maximize mine stability, and maintain the value and reasonable foreseeable use of such surface lands, except in those instances where the mining technology used requires planned subsidence in a predict- able and controlled manner..." Implied in the constraints of Section 516 (b) (1 of SMCRA is a determination of what constitutes "material damage" to surface structures and renewable resource lands, and the capacity for sequential land use following underground mining. This Study was undertaken in an attempt to provide quantitative guidelines for estimating material damage to a wide variety of surface structures and renewable resource lands. Surface structures studied included buildings, bridges, railroads, highways. pipe- lines, d~, embankments, and water reservoirs. Damage criteria for buildings developed in this study along with details of the adopted methodology have been presented elsewhere (Bhattacharya et al.) (1). and will not be repeated here. Similar guidelines for renewable resource lands have also been discussed (Singh and Bhattacharya) (2)

Jan 1, 1984

By Khaled Morsy

In this paper a new method to evaluate the stability of yield pillars was introduced. The evaluation of yield pillar stability was conducted in three steps: (1) Estimation of pillar loading - numerical modeling is a good tool to estimate the pillar loading for most of the geological, geometrical and mining conditions. In this study. finite element technique was used to estimate the state of stress, strain, elastic strain energy and plastic energy, etc. at every point within the pillar. (2) Defining the yield pillar loading zones¬. The proposed method considered the non-uniform stress distribution of yield pillars by dividing the area of the yield pillar into three loading zones, namely core, transition and rib zones. (3) Estimation of stability indicators - an appropriate stability measure was proposed for each loading zone. such as; Core Stability Factor. CSF, Pillar Bump Index, PBI and Rib Instability Index. RIF. A case study of successful and unsuccessful yield pillar application was used to assess the applicability of the proposed method.

Jan 1, 2003

By D. L. Price

This paper presents a method of measuring field power consumption of shearers and deriving specific energy consumption versus shearer haulage speed performance curves. A description of a computer program which simulates the cutting action of a shearer drum is presented. A comparison between the field specific energy curves and the computer simulation results is made. The simulation results compare well with the field data where all the coal seam is being cut by the shearer. The computer simulation results give hack-calculated field pick force cutting characteristics, which allow the performance of the shearer to be predicted for other operating conditions, e.g. effect of changes to pick lacing, effect of changing the drum rotational speed, allows the calculation of power requirements for a particular production rate.

Jan 1, 1992

By Ali Benbia

The paper shows results from the physical modelling of longwall excavations concentrating on the fracture pattern development above longwall panels. The modelling technique is briefly described and the detailed analysis of the fracture pattern is discussed. The physical modelling is particularly useful in visualising the caving process above longwall panels. This advantage has been exploited using the image analysis technique. Black and white pictures have been taken while the models are tested in successive stages and then enhanced and stored as binary images by scanning using an image analyser coupled with a video camera. These images are then processed using a custom C program which allows the discretization and characterisation of the fractures. This provides a valuable data base for further numerical and statistical analysis which subsequently may allow the assessment of the water inflow hazard to the workings. In this paper a sequence of different extraction thickness have been analysed using the technique mentioned above to show the effect of the extraction thickness on the fracture distribution and intensity.

Jan 1, 1993

By Gift Makusha

Total extraction mining of coal is being undertaken on a small scale by Anglo Coal South Africa. Longwall mining has been done in the group at New Denmark colliery since the 80s and recently the company started experimenting with pillar mining at Goedehoop colliery. Total extraction mining methods result in caving of the overburden strata that is often accompanied by displacement of air causing wind blasts. The windblasts can cause injuries to underground mining personnel, damage to mining equipment, and ventilation disruptions. Some incidents of windblast damage and injuries have been recorded at both New Denmark and Goedehoop Colliery that prompted mine management to experiment with the goaf warning device, an instrument developed by CSIR mining technology to give early warning of impending god. The initial version of the Goafwarn was originally developed to operate as a stand-alone unit, with LED lights mounted on the device that would flash if the unit was triggered'. That set up would be effective in small working areas where the flashing lights are visible to all workers. This requirement is more difficult to satisfy in longwall application of the unit where mining activities are spread over large linear distances of up to 250 m and the flashing lights would not be visible to all workers in the longwall face. It was also found that the single unit had limitations in distinguishing between man made vibrations and cracking in rockmass leading to an unacceptably high incidence of false alarms. The case studies describe the experiments that were carried out to ameliorate the two issues mentioned above through networking more than one unit. This was achieved through the development of radio communication from the warning devices to a central control (Trollex Data Acquisition Unit) from which the alarms are issued once the prescribed triggers are satisfied. The operating system was made up of 4 Goafwarn units installed 30 m apart with 2 in the tailgate and other 2 in the main gate. The alarm state was defined where 3 or more units were triggered simultaneously. The concept was applied successfully in real time during longwall mining, and management procedures have been developed to achieve safety through the system including switching off the shearer, and flicking the lights to warn people.

Jan 1, 2005

By Chase K. Barnard, Rahul Thareja, Sean Warren, Rajagopala Kallu

"Inflatable rock bolts are commonly utilized for ground support in Nevada underground mines, however, there is limited information regarding what factors influence the bond strength of these bolts. Bond strength is an important parameter in friction bolt support design, and this information is usually not available from manufacturers as a standard. Previous research has investigated the bond strength of friction bolts; however, these studies focused primarily on split set bolts and ground conditions more favorable compared to those commonly encountered in Nevada underground mines. Without site-specific bolt pull-out tests, ground control engineers are required to make assumptions regarding the in-situ bond strength of rock bolts. This is especially a concern for projects in the feasibility and initial development stages before site-specific experience is acquired.The purpose of this paper is to establish confidence in anticipated minimum bond strength for inflatable rock bolts by comparing the bond strength to variable geotechnical conditions using the Rock Mass Rating (RMR) system. To investigate a correlation between these parameters, the minimum bond strength of pull-out tested inflatable rock bolts was compared to the RMR of the rock in which these bolts were placed. Bond strength vs. RMR plots indicate that expected minimum bond strength is positively correlated with RMR, however the correlation is not strong. Cumulative percent graphs indicate that 97% of pull-out tests result in a minimum bond strength of 1 ton/ft and ½ ton/ft in RMR =45 and <45, respectively.Although lower bond strengths are more commonly encountered in low RMR ground, high bond strengths are possible as well, yielding higher variability in bond strengths in low RMR ground. Bond strength of friction bolts relies on contact between the rock bolt and drill hole. Experience in Nevada indicates that RMR is known to affect both the quality and consistency of drill holes which likely affects bond strength. Drilling and bolting in low RMR ground is more sensitive to drilling and bolting practices, and strategies for maximizing bond strength in these conditions are discussed."

Jan 1, 2015