Main Sections of Report
NIOSH Activities Related to the Geotechnical Project
Cape Breton Development Corporation Prince Mine
Quinsam Coal Corporation - 2 North and 4 South Mines
Smoky River Coal Corporation 5B-4 Mine
Conclusions and Baseline Comparison
Appendix A. The Coal Mine Roof Rating (CMRR)
Appendix B. Mine Design for Control of Horizontal Stress
Appendix C. The Analysis of Longwall Pillar Stability (ALPS)
Appendix D. Analysis of Retreat Mining Pillar Stability (ARMPS)
Appendix E. Calculation of the CMRR for the Canadian Mines
List of Tables
Table 1 - Comparison of Entry Stabilty Issues at Prince and Phalen Mines
Table 2 - Baseline Comparison of Three Canadian Mines
List of Figures
All figures are located in Appendices A to E
A1. Relationship between CMRR and entry width for U.S. longwall mines.
A3. Relationship between CMRR and the stability of extended cuts.
B1. Horizontal stressfields in the U.S.
B2. Concentration of horizontal stress around longwall gob areas.
C1. Loads on longwall pillars estimated by the ALPS program.
C2. Design equation for U.S. longwalls using ALPS and the CMRR, derived from case histories.
C3. Australian ALTS design equation and case histories.
D1. Features of retreat mining layouts considered in the ARMPS program.
D2. Room-and-pillar case histories in the ARMPS data base.
CMRR Rating Sheets, Unit Rating Sheets, Lithology-Borehole Logging Sheets and Memo
The National Institute for Occupational Safety and Health (NIOSH) and CANMET Mining and Mineral Sciences Laboratories have undertaken a cooperative agreement regarding the Underground Coal Mining Safety Research Collaboration (UCMSRC). The UCMSRC was initiated by CANMET as a collaborative 'in-kind consortium' which brings together all of the principal stakeholders in Canadian underground coal mines; including industry, labour, regulators, inspectors, and universities. Currently, there are three active underground coal mines in Canada:
The Agreement calls for NIOSH to participate in a UCMSRC project titled "Standardization of Geological and Geomechanical Assessment at Underground Coal Mines in Canada." The NIOSH contribution was to provide a senior ground control specialist (Dr. Christopher Mark) to visit each of the three mines. Each visit was to include:
At the conclusion of the project, a report was to be prepared outlining current geological and geomechanical assessment and application/design practice, a comparative analysis, and a recommended "baseline" which operators, regulatory agencies, and mineworkers can use at all sites. In addition, an outside publication on the lessons learned from the collaboration will be prepared and presented to U.S. and Canadian mining audiences.
This is the final report called for in the Agreement. It begins with a brief summary of NIOSH activities on the Project, followed by individual reports on each of the three mines. Conclusions and a baseline comparison are then summarized. Appendices contain the CMRR and other rock properties data that were collected through underground assessments and rock testing, and brief descriptions of the CMRR, horizontal stress, ALPS and ARMPS.
Dr. Christopher Mark made three trips to Canada to complete the NIOSH activities under the Project. The first, to Cape Breton, NS, took place during October 27-29, 1998. Activities included:
The second visit was to Grande Cache, AB, during the week of April 27-May 1, 1999. Activities included:
While in Grande Cache, Dr. Mark also attended and gave a presentation to the UCMSRC Executive Committee and Technical Forum.
The final visit, to Campbell River, BC, took place from November 8-12, 1999. Activities there included:
The Prince Mine uses the longwall method of coal extraction, with Dosco dintheaders and roadheaders for gate entry and mains development. Prince was opened in 1974, and is today the last mine operating in the Sydney coalfield. Currently, the longwall is being moved to the first panel of a new longwall block to the North of previous workings.
Pillar Design
The depth of cover is about 320m, and dips are less than 5 degrees in the current North longwall reserve. There is no previous mining above or below Prince Mine. Two gate entries are driven for each longwall, separated by a 60m wide abutment pillar. There are typically no more than two crosscuts (cundies) per panel. Entries are nominally 4.5m wide, but often spall to 5m. The Hub seam averages 2.3m in thickness.
The ALPS stability factor for recent Prince longwalls is approximately 2.3 for tailgate loading. This may be compared to the 1.34 suggested by the U.S. case history data assuming a CMRR of 30. The Australian ALTS equation suggests that an SF of 1.8 should be adequate.
Horizontal Stress
While there have been no recent in situ stress measurements at Prince, every indication is that the horizontal stresses are similar to those found elsewhere in the Sydney coalfields. At Phalen, measurements found that the maximum horizontal stress was oriented N60E, and was approximately 2.5 times as great as the vertical stress.
Most of the longwall panels at Prince have been oriented N70W, so that they intercepted the maximum horizontal stress at an angle of about 50 degrees. Conditions on development were very poor, with significant movement in both the roof and the floor. Often a row of center posts had to be installed within 60m of the development face, as I observed on 15 West. Fortunately, the headgate was shielded from a horizontal stress abutment.
The most recent panel, 9 East, was the first to oriented N45E, or just 15 degrees off parallel with the maximum horizontal stress. The development conditions I observed there were much improved. Moreover, the headgate in 9 East and in the similarly-aligned North panels should be in a stress shadow.
Coal Mine Roof Rating
The immediate roof was observed to be an underclay, heavily rooted, with closely spaced bedding. It was also moisture sensitive. Its Unit Rating was about 30, before exposure to moisture, and 23 after exposure. The thickness of this unit was 1.5m and 1.8m in two cores I evaluated. In future longwalls, it may be as much as 6m thick. In my underground observations, a slightly stronger 0.8 m silty shale was present 0.8 m into the roof.
The underclay is overlain by a medium grained, grey, medium bedded (10 cm spacing) sandstone. The sandstone's strength is less than most sandstones (about 50 MPa). In one core it contained coal spars, in another, it was interbedded with thin shale bands. Its Unit Rating is in the 45-55 range. I did not observe the sandstone underground.
When the underclay is less than 2m thick, 2.6 m bolts could be used and the overall CMRR would be about 40. As the underclay thickens, the CMRR reduces to a Unit Rating of 30 and even lower, when the long-term effect of moisture is considered.
Overall, the CMRR>30 roof at Prince is very weak in comparison with most U.S. or Australian mines.
Roof Support
The primary roof support at Prince is provided almost exclusively by steel sets. The supports are W6*25 beams with W6*15 legs. Their capacity is 12 tonnes yield and 18 tonnes ultimate for a 4.3 m span (assuming an evenly distributed load and a 45 degree load triangle). When the mine was first opened, in the 1970's, fully-grouted resin bolts were used in a small area that was mined by room-and-pillar. Reportedly, the bolts worked well when they anchored in a sandstone channel, but several roof falls occurred beneath thick shale roof in the transition zone.
The two-entry longwall system used at Prince is very intolerant of roof falls. To control the roof, the entries need to be as narrow as possible. It is fortunate that there are so few intersections (and no 4-way intersections). Extended cuts are out of the question, and place-changing would not be recommended. Roof bolts should be installed as soon as possible after mining, as close as possible to the face.
Long roof bolts could possibly be used as sole support where the underclay was less than 2m thick and the sandstone was at least 1m thick. They would be employed with a surface support (like mesh) to prevent the weak immediate roof from unraveling. Where the sandstone was higher up or thinner, supplemental support (cable bolts or trusses) would probably be required. The specific type and pattern of supplemental support would depend on the location, thickness, and strength of the sandstone. However, the overall capacity would probably not have to greatly exceed the 18 tonnes/m of the current steel supports. Roof monitoring would be very important to locate any areas of aggressive roof that required tertiary support. Given the height and narrowness of the entries, pattern rib support should probably be considered as well.
While roof conditions are generally more difficult than encountered in most U.S. mines, entry stability should be easier to maintain at Prince than it was at Phalen. Table 1 compares attributes of the Prince 1 North panel headgate to the Phalen 8 East Bottom where a major roof fall occurred. Four factors (depth, multiple seam interactions, headgate horizontal stress concentration, and entry width) are more favorable at Prince. Only one, the CMRR, is somewhat less favorable.
| Issue | Phalen 8E Bottom | Prince 1N |
|---|---|---|
| Depth of Cover | 680 m | 320m |
| Multiple Seam? | YES | NO |
| Orientation for Horizontal stress | Good | Good |
| Horizontal stress concentration? | YES | NO |
| Entry Width* | 4.9 - 5.5 | 4.5 - 5 m |
| CMRR | 35 | 30 |
Quinsam Coal Corporation has operated underground mines on Vancouver Island since 1991, following several years of surface mining. Production has come exclusively from room-and-pillar mining using continuous miners and shuttle cars. The current active underground operation is the 2 North Mine in the 1 Seam. The 4 South Mine in the 3 Seam is currently on standby.
Pillar Design
The 1 Seam is 4-5 m thick, and dips approximately 12% (7 deg) in the current mining area. Entries are mined 6 m wide by 3 m high, with floor coal recovered during depillaring. At the time of my visit, pillar recovery was being conducted in 7 and 8 Panels, and 9 Panel was being developed. The pillars in the panels are recovered using a Christmas-tree cut sequence.
The entries in 8 and 9 Panels were developed on 16 m centers, with crosscuts on 36 m centers. The pillars in 7 panel were originally developed for a split-and-fender mining system, and measured approximately 32 m square.
Cover is relatively light at Quinsam Coal, and is reportedly not more than 140 m at the 2 North Mine. The Analysis of Retreat Mining Pillar Stability (ARMPS) software developed by NIOSH (see Appendix D) is used as the pillar design guideline at Quinsam. The ARMPS Stability Factor is well over 2.0 for all phases of the current mining plan at 2 North.
In 4 South, concerns about surface subsidence (under a swampy area) and controlling the caving of the massive sandstone roof have precluded full pillar extraction. Originally the mine practiced partial pillar recovery with recovery pillars ‘in line’. However, because of localized roof falls which occurred as a result of having the partially depillared areas in line, a "checkerboard" partial pillar pattern was adopted instead. The checkerboard plan successfully mitigated the degree of localized roof falls.
Horizontal Stress
No evidence of horizontal stress was observed at either mine.
Geology and Coal Mine Roof Rating
Roof geology observations were made at an overcast near the mouth of 9 Panel in 2 North Mine, and at a fault at the bottom of the mains in 4 South. In addition, roof cores were obtained from two locations in 2 North and one in 4 South. All core was taken in close proximity to current working areas.
In general, the 1 seam is overlain by grey shale, which grades into a slightly more competent silty shale about 0.6 m up in the roof. The structural competence of the units are reduced in periodic fracture zones which affect 10-15% of the roof. The fractures tend to follow the strike, which means that entries driven on strike tend to be somewhat less stable than those driven with the dip. At 2 North Mine, the mains have been driven along strike, and panels are driven down dip.
The difficulty of driving panels through these faults and supporting the ground in these areas has resulted in the reduction of the number of entries as the panels pass through the fault. The faults result in lost reserves, difficult ventilation and uneven cave lines when depillaring. The faults would be expected to cause even greater difficulties if the panels were driven parallel with them. This situation is illustrated by the offsets of the supply road of 4 Mains that were necessitated by the fault which runs along the direction of the roadway. There was some discussion at the mine of changing to ‘strike mining’ off 9 Panel. Although mining along strike might be preferable from an operations standpoint, since the faults also tend to parallel the strike, ground conditions would probably be less favorable. In addition, it is not likely that pillar stability would benefit from mining along strike, because recovering pillars up dip allows some portion of the weight to be transferred toward the gob.
Underground, the strength of the shale was judged to be about 70 MPa. Its bedding was closely spaced, but the rock did not fail easily on bedding, instead preferring to create rough fractures through intact material. The overall CMRR was calculated at 54 using the underground data, and this would be reduced by 4 points in the fracture zones.
The RQD of the lower mudstone was about 60, with a fracture spacing of 10 cm. The siltier, upper mudstone’s RQD was about 90 with a fracture spacing of 30 cm.
Point Load Tests of the lower mudstone found an axial strength of 20-40 MPa, and a diametral strength of about 6 MPa. The axial strength of the upper mudstone was 44-64 MPa, which compared well with the uniaxial compressive strength of 57 MPa. The diametral strength of the upper mudstone was about 12 MPa.
Unit Ratings were estimated at 38 for the lower mudstone and 51 for the upper. The overall CMRR was estimated at 47 from the core data.
A study of experience with extended cuts at 36 U.S. mines found that when the CMRR was greater than 50, cuts deeper than 6 m could almost always be achieved. When the CMRR was less than 38, extended cuts were almost never stable. However, extended cut lengths in the U.S. can be limited by dust control and ventilation issues, particularly where the mine does not employ a blowing face ventilation system with scrubbers on the miner.
At Quinsam, the mine is only in production 1 shift per day. This makes bolting all cuts within the same shift extremely difficult and within 24 hours very difficult on weekends. While it is good mining practice to bolt the roof as soon as possible after it is mined, minimizing the time lag is most important in weaker ground. At Quinsam, every effort should be made to bolt roof that is affected by faults or fracture zones quickly.
In 4 South, the roof is a medium-grained, massive sandstone with only occasional joints. Its strength was estimated at 70-130 MPa underground. Uniaxial compressive strength and axial point load tests found a strength of 55-70 MPa. The diametral strength was about 30 MPa. Overall, the CMRR was estimated at 78 from underground observations and 68 using core data.
Most of the layers within the 1 Seam in 2N mine seemed to be relatively hard, dull coal. The roof contact is distinct but very strong and shows no tendency to slip. About 2 m below the roof there is a very soft mud seam whose thickness varies over the property. In general, the ribs were almost entirely intact, even at pillar corners.
Roof Support
Roof support at 2 North Mine is provided by 2.4 m, 19 mm, Grade 60, torque-tension bolts installed with a 0.6 m of resin anchor resulting in almost 0 .9 m of encapsulation. The break out of the resin mix is achieved with a pinched thread on the roof bolt. The pattern is to install a row of 5 bolts every 1.2 m. In the U.S., a 5 bolt pattern would be very unusual where a dual-boom roof bolter is used. Welded wire mesh panels are installed to support any spalling of the lower mudstone that might occur over time due to weathering or due to changing stresses during pillaring. Reportedly, no unplanned roof falls have ever occurred except during pillaring and in fracture zones. I thought the "gripper plates" that are used to install the mesh in outby areas was an excellent idea.
Originally, 1.8 m roof bolts were used at 2 North Mine, with 2.4 m bolts reserved for bad ground. Following the fatal accident in 1998, all bolts are 2.4 m long. When breaching faults, entry widths are reduced from the nominal 6 m to 5.2-5.7 m.
Roof support at 4 South is achieved with 1.8 m roof bolts identical to the bolts in 2 North in every other way. The pattern is to install a row of 5 bolts every 1.2 m. No mesh is used due to the competency of the roof and its resistance to weathering. In general, roof of this competency would be supported with shorter mechanical shell or fully-grouted untensioned resin bolts in the US.
Retreat Mining
Retreat mining, including pillar recovery and floor mining, is integral to the mining plan at Quinsam Coal. Since the fatal accident that occurred two years ago, Quinsam has paid close attention to pillaring procedures. I was impressed with the way that pillaring cut sequences are posted for all crew members to see. When I was there, 8 Panel was recovering pillars in a fracture zones, and I was also impressed to see that the plans were adjusted to leave fenders between cuts and to leave some areas completely. Marking the known fracture traces on the roof and on the maps is also an excellent practice.
In my opinion, the key to successful pillar recovery is to support the intersection. Because of the thick seam and the floor recovery at Quinsam, mobile roof supports are not used, and timbers are not used to the degree that they probably would be in the US. Warning timbers were being installed at each cut as a visual indication of roof behavior in the gob, and breaker posts were installed at intersections and in crosscuts to prevent gob over ride. Due to the fewer number of timber used relative to the US, the final stump at the intersection is even more critical than ever. The plans I saw called for a substantial final stump to be left when pillar recovery is complete. I completely endorse this practice. In my experience, a stump about 3 m square should provide effective intersection support without being so strong that it seriously inhibits the roof caving.
Smoky River Coal Corporation has operated underground mines in the foothills of the Canadian Rockies since 1969. While longwalls have been tried on several occasions, most production has come from room-and-pillar mining using continuous miners and shuttle cars. Mining has been conducted in the 4 seam in both the Smoky River and Sheep Creek valleys. The current underground operation, the 5B-4 mine, is located in the Smoky River valley slightly to the north of the No. 5-4 Mine, which contains some of the earliest workings on the property.
Historically, roof conditions have always been more difficult in the Smoky River valley. The original No. 5-4 Mine was reportedly "closed within a short period of time due to strata problems" (Grant et al., 1978). Recent mining in the 9G, 9H, and LB4 mines were much less troublesome.
Pillar Design
At 5B-4 Mine, the reserve is bounded by steeply -dipping thrusted zones to the north and south. At the time of the visit, a 5-entry mains was being driven to the east, with 4-entry panels driven up-dip on the right-hand side. The pillars in the panels were being recovered using a Christmas-tree cut sequence.
The seam in the 5B-4 Mine is 3-5 m thick. Entries are initially mined 5 m wide by 3 m high, with floor coal recovered during depillaring.
The mining depth for the 4 South Panel was about 275 m, but much of the 5B-4 reserve is under as much as 550 m of cover. None of the previous Smoky River mines have been more than 300 m deep (Cain, 1999).
Dr. Peter Cain has carried out extensive analyses of past pillar performance in the 4 Seam using the Analysis of Retreat Mining Pillar Stability program (ARMPS). He has reported his work elsewhere (Cain, 1999).
Horizontal Stress
The Smoky River area has clearly been subjected to intense horizontal tectonic pressures, as evidenced by the pervasive thrust faulting. According to data published by the World Stress Map Project (Zoback, 1992), the deep borehole breakouts show that the eastern Canadian Rockies are currently subject to the major regional North American stressfield, with the maximum horizontal stress oriented approximately N60E (figure B1). One underground observation at the No. 2 Mine reported by Jeremic (1985) indicated that the maximum stress was oriented approximately N35E.
During our visit, there was relatively little evidence of horizontal stress underground. In one location near the outcrop, a few large roof "pots" were indicative of a N30E horizontal stress. It is possible that as mining progresses into the ridge towards deeper cover, the horizontal stress may increase and rotate towards the regional stress orientation.
Coal Mine Roof Rating
Roof geology observations were made at the abandoned 9A Mine highwall in the Sheep Creek valley, and at underground roof falls in the 5B-4 Mine. In general, the 4 seam is overlain by shale or carbonaceous shale, which is in turn overlain by a stronger siltstone or sandstone. The structural competence of the units are reduced by periodic fracture zones related to thrust faults.
At the highwall, the carbonaceous mudstone was 0.6-1.0 m thick. It had a strength of 35-75 MPa, was highly laminated, and typically contained joints on a 0.6-1.6 m spacing. Within fracture zones, the joint spacing was decreased to less than 0.2 m. Its Unit Rating was 35-40. The strength of the bedded, "wispy" siltstone appeared to exceed 120 MPa. Its Unit Rating was typically 65, reduced to 55 within fracture zones. Overall, the CMRR ranged between 57 (normal) and 48 (in fracture zones).
The rock types were similar in the 5B-4 mine, except that:
The overall CMRR was estimated at 50 for "normal" conditions, and 40 within fracture zones.
A study of experience with extended cuts at 36 U.S. mines found that when the CMRR was greater than 50, cuts deeper than 6 m could almost always be achieved. When the CMRR was less than 38, extended cuts were almost never stable. This pattern seems to hold for Smoky River, where extended cuts have reportedly been successful everywhere but in fracture zones in the 5B-4 and other Smoky River mines.
Peter Cain provided some rock mechanics test data on core from the 5B-4 mine. Using the most recent conversion factors for Point Load Test data, the strength values are:
| Mudstone: | Diametral PLT strength = 0.33 MPa | Axial PLT strength = 28.5 MPa |
| Grey Siltstone: | Diametral PLT strength = 24.6 MPa | Axial PLT strength =103 MPa |
Using these data and the CMRR-for-core procedures, the Unit Ratings are estimated at 40 for the mudstone and 76 for the siltstone. The latter would be reduced, however, if the fracture spacing of the siltstone core was less than 2 m.
Most of the layers within the coal seam were relatively hard, but there were at least 5 zones of soft, platy, sheared coal in the top 4.3 m. A particularly weak zone occurred at the top of seam just below the roof shale. These zones facilitate outward movement of the ribs under load.
Roof Support
Mechanical roof bolts provided support in the early days of the Smoky River mines. Numerous roof falls occurred in the No. 5 and No. 2 mines, and by 1975 resin bolt trials were initiated (Grant et al., 1978). Conditions improved when mining shifted to the Sheep Creek valley. There rows of 4 to 6 mechanical bolts, 2.4 m long, usually provided adequate support for the 6.7 m wide entries (Cain, 1998).
In the 5B-4 mine, entry widths have been reduced to 4.8 m (nominal). The Manager’s Support Rules for main entries call for fully grouted resin roof bolts, 3.6 m long, installed in rows of 4 every 1.2 m, with steel straps and mesh. In addition, one or two 1.5 m mechanical rib bolts are installed per row, also with mesh. Support is similar in the short-term roadways, except that 2.4 m mechanical bolts are used in place of the 3.6 m resin bolts.
Roof stability monitoring is provided by 2-point telltales installed approximately every 60 m.
By U.S. standards, the level of support at 5B-4 is relatively heavy. I am not aware of any mines that use 3.6 m long bolts routinely. With a similar CMRR and entry width, 2.4 m bolts would be more typical. Longer bolts might be used in intersections.
Currently, the same support system is employed in both fracture zones and unfractured ground. The unfractured roof probably does need the same level of support, however. If the boundaries of the fracture zones could be accurately identified by the bolting crew, it might be possible to use a lower level of support elsewhere.
During the bolting process, I noticed that the miners needed to go beyond the last row of bolts to place the screens above the ATRS. I have since found that Fletcher (the roof bolt machine manufacturer) makes screen-holding devices, which are attached to the ATRS and eliminate the need to expose workers to unsupported roof.
I thought the system of plastic tubes that the miners used to insert the resin was very ingenious. I also observed that the resin seemed to take an unusually long time to set. It seems likely that a resin manufacturer could provide a faster resin that was suited to the conditions at Smoky River.
Retreat Mining
Retreat mining, including pillar recovery and floor mining, is integral to the mining plan at Smoky River.
Pillar recovery is always hazardous, but under deep cover it appears it becomes even more risky. Analysis of recent accidents in U.S. mines shows that a disproportionate share of fatalities have occurred in deep cover pillaring operations. During 1998 alone, there were fatal accidents in West Virginia, Kentucky, and Colorado (the last was a double fatality). In each case the cover exceeded 200 m. It is important that every miner know and follow standard pillar recovery procedures to limit their exposure to hazards. In the U.S., shuttle cars used for pillar recovery (and development for that matter) would be equipped with canopies.
References
Cain P (1999). Developments in Coal Pillar Design at Smoky River Coal Limited, Alberta, Canada. Proc. 2nd Int. Workshop on Coal Pillar Mechanics and Design, NIOSH IC 9448, pp. 15-22.
Cain P (1998). Experience with the First Modern Longwall Installation in Western Canada at Smoky River Coal Limited, Grande Cache, Alberta.
Grant F, Wigelsworth RM, Charlton K (1978). Underground Strata Control with Resin Grouted Roof Bolts in McIntyre Mine’s Coal Operations. Proc. First Intl. Symp. On Stability in Mining, Vancouver, pp. 197-213.
Jeremic, ML. Strata Mechanics in Coal Mining. Balkema, Rotterdam, p. 149.
Zoback, ML (1992). Stress Field Constraints on Intraplate Seismicity in Eastern North America. J. Geophys. Res., v. 97, no. B8, pp. 11,761-11,782.
The three Canadian underground coal mines display a wide variety of geologies, mining methods, and support techniques. The Prince Mine, the only longwall of the three, generally faces the most difficult ground control conditions. Prince has the weakest roof, the greatest level of horizontal stress, and usually the greatest cover.
Quinsam 2 North Mine, at the other extreme, has the most benign conditions. It is the shallowest of the 3 mines, has the strongest roof, and shows no evidence of horizontal stress. It is not surprising that Quinsam uses the least support, and the widest entries, of the three mines.
The 5B-4 Mine at Smoky River falls between the two end points. Fracture zones are common at this mine, reducing the CMRR to near 40. As a result, narrow entries and a high level of roof support are required. Conditions promise to get even tougher as the depth of cover increases. A decision has now been made to recover 5B-4 Mine, and replace it with a new mine in a different part of the reserve. If conditions at the new mine resemble the ones encountered historically in the Sheep Creek valley, they will be comparable to those at Quinsam.
Table 2 sums up the baseline comparison for the three mines.
Many of the ground control safety technologies developed in the US seem to have direct application to Canadian mines. Pillar design and the CMRR seem particularly relevant. There are some aspects of Canadian underground coal mining that seem to be unique. Research in these could be very helpful for the future of the industry. The two most important might be:
| Prince | Smoky River 5B-4 | 2 North Quinsam | |
|---|---|---|---|
| Development mining method |
Single-entry Roadheader |
Room-and-pillar Continuous miner |
Room-and-pillar Continuous miner |
| Retreat mining method |
Longwall | Full pillar extraction with floor recovery |
Full pillar extraction with floor recovery |
| Depth of cover | |||
| Seam thickness | |||
| Dip | |||
| Pillar width | |||
| Entry width | |||
| Pillar SF | |||
| Horizontal stress | |||
| CMRR | |||
| Primary roof support |
Steel arches | 3.6 m fully grouted roof bolts, with steel straps and mesh |
2.4 m torque tension bolts, occasional mesh |
| Rib support | Steel arches | 2-1.5 bolts with mesh | None |
| Major ground control problems |
Weak roof Horizontal stress Floor heave |
Fracture zones Excessive dips Pillar recovery |
Faults/fracture zones Pillar recovery |
One reason that roof falls have proven so stubborn a problem is that mines are not built of man-made materials like steel or concrete, but rather of rock just as nature made it. The structural integrity of coal mine roof is greatly affected by natural weaknesses including cracks, small faults, and layering. To make matters more difficult, the geologic processes that formed it varied in space and time, so engineering properties of the roof can change dramatically from mine to mine (and even within individual mines!).
Engineers have had difficulty obtaining quantitative data on the strength of rock masses for design. Traditional geologic reports contain valuable descriptive information, but seldom include engineering properties. On the other hand, laboratory strength test results are inadequate because the strength of a small rock sample is only indirectly related to the strength of the rock mass.
To help quantify the engineering properties of mine roof, the Coal Mine Roof Rating (CMRR) was proposed (Molinda and Mark, 1994). The CMRR weighs the geotechnical factors that determine roof competence, and combines them into a single rating on a scale from 0 to 100. The CMRR integrated 20 years of research on geologic hazards in mining with worldwide experience with rock mass classification systems. In developing the CMRR, field data were collected from nearly 100 mines in every major U.S. coalfield. Cost-sharing cooperative research agreements were signed with the Cyprus, Ziegler, and Peabody coal companies to support the research.
The CMRR can be calculated from underground exposures like roof falls and overcasts (Molinda and Mark, 1994), or from exploratory drill core (Mark and Molinda, 1996). In either case, the main parameters that are measured are:
Simple index tests and observations are used to rate each of these parameters. In the case of drill core, Point Load tests (PLT) are used to estimate the compressive strength and the cohesion. The CMRR is obtained by summing the ratings for the individual parameters. Detailed procedures for calculating the CMRR have been published elsewhere (Molinda and Mark, 1994; Mark and Molinda, 1996). A computer program is currently being developed to aid in the collection, interpretation, and presentation of CMRR data.
The CMRR makes three significant contributions:
The CMRR has found numerous applications in ground control design. A study conducted at 44 longwall mines found that tailgate performance was largely determined by the CMRR and the ALPS pillar stability factor (Mark and Chase, 1994). Significant correlations between the CMRR and both entry width (figure A1) and the intensity of roof support were also found. Another study determined that yielding pillar gate entry designs have only been successful when the CMRR was greater than 50 and the pillar’s width-to-height ratio was less than 5 (DeMarco, 1994).
Figure A1 - Relationship between CMRR and entry width for U.S. longwall mines.
Data has also been presented that relate the incidence of roof falls to the CMRR and intersection span (figure A2). These were based on observations at five underground mines (Mark et al., 1994). The CMRR has lately been incorporated into guidelines for multiple seam mine design (Luo et al., 1997), hazard analysis and mapping (Wuest et al., 1996), tailgate support selection (Harwood et al., 1996), and feasibility studies (Beerkircher, 1994). The Mine Safety and Health Administration has used the CMRR in fatal accident investigations, and at least three major coal companies have recently taken steps to integrate the CMRR into their exploration programs.
A recent study (Mark, 1999) applied the CMRR to extended (deep) cut mining, where the continuous miner advances the face more than 20 ft beyond the last row of permanent supports. Data on the CMRR and extended cut experience were collected at 36 mines in 7 states. It was found that when the CMRR was greater than 55, deep cuts were routine in nearly every case. When the CMRR was less than 37, extended cuts were almost never taken. Between 38 and 55, extended cuts were feasible sometimes but not others (figure A3). The data also show that extended cuts are less likely to be stable if either the entry span or the depth of cover is increased.
Figure A3 - Relationship between CMRR and the stability of extended cuts.
References
Beerkircher MD (1994). Monterey Coal Company’s Longwall Project. Proc. IL Mining Institute., Collinsville, IL, pp. 85-93.
DeMarco, MJ (1984). Yielding Pillar Gateroad Design Considerations for Longwall Mining. Paper in New Technology for Longwall Ground Control: Proceedings of the USBM Technology Transfer Seminar, USBM SP 94-01, pp. 19-36.
Harwood, C., M. Karmis, C. Haycocks, and J. Luo (1996). Optimizing Secondary Tailgate Support Selection. Proc. 15th Intl. Conf. on Ground Control in Mining, Golden, CO, pp. 469-476.
Luo, J., C. Haycocks, and M. Karmis (1997). Gate Road Design in Overlying Multiple Seam Mines. SME Preprint 97-107, 11 pp.
Mark, C. (1999). Application of the Coal Mine Roof Rating (CMRR) to Extended Cuts. Mining Engineering, April, pp. 52-56.
Mark, C., F. E. Chase (1994). Design of Longwall Gate Entry Systems Using Roof Classification. Paper in New Technology for Longwall Ground Control: Proceedings of the USBM Technology Transfer Seminar, USBM SP 94-01, pp. 5-18.
Mark, C., Molinda GM, Schissler AP, and Wuest WJ (1994). Evaluating Roof Control in Underground Coal Mines Using the Coal Mine Roof Rating. Paper in the Proceedings of the 13th Conference on Ground Control in Mining, pp. 252-260.
Mark, C. and G. M. Molinda (1996). Rating Coal Mine Roof Strength from Exploratory Drill Core. Proceedings of the 15th International Conference on Ground Control, Golden, CO., pp. 415-428.
Molinda, G. and C. Mark (1994). The Coal Mine Roof Rating (CMRR)--A Practical Rock Mass Classification for Coal Mines. USBM IC 9387, 83 pp.
Molinda, G. M., and C. Mark (1996). Rating the Strength of Coal Mine Roof Rocks. USBM IC 9444, 36 pp.
During the past 15 years, horizontal stress has become central to an understanding of coal mine ground control. An important breakthrough was the recognition that the stresses observed in coal mines are caused by global plate tectonic forces (Mark, 1991). The World Stress Map Project (Zoback and Zoback, 1989) has identified stress regimes in many parts of the world, by analyzing active faults, borehole breakouts, and hydraulic fracturing stress measurements. In eastern North America, their data shows that the maximum horizontal direction is typically oriented East-Northeast (Figure B1).
An evaluation of stress measurements made in underground coal mines confirmed that stress direction in mines (Mark and Mucho, 1994). The measurements also showed that the magnitude of the maximum horizontal stresses is typically three times greater than the vertical. The horizontal stressfield is biaxial, with the maximum horizontal stress usually about 40% greater than the minimum. Near the earth's surface, however, topographic features can effect both the orientation and the magnitude of the horizontal stress.
Several other factors also determine the degree to which horizontal stress will affect ground control:
Roof type- Weak roof rock is more likely to suffer damage than strong rock, and laminations greatly reduce a roof's ability to resist horizontal stress.
Entry orientation - Entries that are aligned with the maximum horizontal stress will suffer less damage on development than those perpendicular to it.
Longwall orientation- Horizontal stress cannot pass through a gob area, so zones of stress relief and stress concentration are created (figure B2). Their location depends on the panel orientation, the direction of retreat, and the sequence of longwall panel extraction.
Stream valleys - Stream valleys are often associated with difficult horizontal stress conditions. Stream valleys can also change the orientation of the maximum stress (Molinda et al., 1991).
Figure B2 - Concentration of horizontal stress around longwall gob areas.
Stress measurements are too expensive for most mines to use routinely. As a substitute, stress mapping procedures have been developed to estimate the direction of the maximum horizontal stress (Mucho and Mark, 1994). Such features as roof "guttering" or roof "pots" are observed underground, and the stress direction is inferred from their orientation and severity.
Horizontal stress concentrations are most likely to occur in the headgate, or in the tailgate of a first panel. After one panel has been completed, horizontal stresses are largely relieved in the tailgate. A recent paper (Mark et al., 1998) concluded that proper panel orientation and sequence is the key to maintaining headgate ground control. The optimum orientation is not parallel with the maximum horizontal stress, as previously thought, but rather 20° in the stress shadow of the gob. Stress concentrations can also occur near pillar extraction workings.
References
Chase, FE, Mark C, Mucho TP, Campbell, PL, and Holbrook, AD [1999]. The Advance and Relieve Mining Method: A Horizontal Stress Control Technique. Proc. 18th International Conference on Ground Control in Mining, Morgantown, pp. 300-308.
Mark C[1991]. Horizontal Stress and Its Effects on Longwall Ground Control. Mining Engineering, Nov., pp. 1356-1360.
Mark C, Mucho TP[1994]. Longwall Mine Design for Control of Horizontal Stress. Paper in New Technology for Longwall Ground Control: Proceedings of the USBM Technology Transfer Seminar, USBM SP 94-01, pp. 53-76.
Mark, C., Mucho TP, and Dolinar D (1998). Horizontal Stress and Longwall Headgate Ground Control. Mining Engineering, January, pp. 61-68.
Molinda GM, Heasley KA, Oyler DC, Jones JR[1991]. Effects of Surface Topography on the Stability of Coal Mine Openings. Proc. 10th Intl. Conf. on Ground Control in Mining, Morgantown, pp. 151-160.
Mucho TP, Mark C[1994]. Determining the Horizontal Stress Direction Using the Stress Mapping Technique. Proc. 13th Intl. Conf. on Ground Control in Mining, Morgantown, pp. 277-289.
Zoback, ML, Zoback, MD[1989]. Tectonic Stressfield of the United States. Ch. in Geophysical Framework of the Continental U.S., Geol. Soc. Am. Memoir 172, pp. 523-539.
Longwall pillars perform the essential function of protecting the gate entries which provide the only access to the face. Blockages in the gates disrupt ventilation patterns and close off emergency travelways, posing serious safety hazards. In addition, the impact of any downtime is magnified on a longwall because such a high proportion of the total mine productive capacity is concentrated at the longwall face.
The Analysis of Longwall Pillar Stability (ALPS) method was developed in 1986 to answer the need for effective longwall pillar design. The fundamental premise of the ALPS method is that there is a strong correlation between pillar performance and gate entry stability. Specifically, it has been found that if the pillars maintain a proper load-bearing capacity relative to the applied load, then good mining conditions can be maintained. The method has three components:
ALPS divides the loads applied to longwall pillars into two parts. These are development loads, which are present before longwall mining, and abutment loads, which arrive during longwall panel extraction (Figure C1).
To estimate the load-bearing capacity of the longwall pillars, ALPS uses the Bieniawski formula. The load-bearing capacity of a longwall pillar system is then calculated as the sum of the individual pillar resistances. The ALPS SF is defined as the total load-bearing capacity of the longwall pillar system divided by the total applied load.
Pillar design is not the only factor that affects longwall gate entry performance, however. For example, a mine with strong roof rock will have better conditions, all other things being equal, than a mine where the roof is weak. The entry width and the amount of artificial support can also be important.
Studies were conducted at 44 U.S. longwall mines representing all the major U.S. coalfields. At each mine, underground observations of site geology, entry conditions, and support design were recorded. Discussions with mine personnel eventually identified 69 longwall case histories, with each gate entry design characterized as satisfactory, unsatisfactory, or borderline. The CMRR was used to define the roof rock in longwall case history data base, and the ALPS stability factor (SF) was used to define pillar design. Other rating scales were developed for primary support, secondary support, and entry width.
Statistical analyses indicated that a simple equation could be used to guide the design of longwall pillars and gate entries:
ALPS SFR = 1.76 - 0.014 CMRR
Where: ALPS SFR = ALPS SF suggested for design.
The design equation is illustrated in Figure C2.
Most recently, a research project was funded by the Australian Coal Association Research Program to calibrate ALPS for Australian conditions (Colwell et al., 1999). There were some geotechnical and mine layout differences between United States and Australian coalfields, including the use of two-entry gates and longer spacing between crosscuts, that required investigation. Ultimately, case history data were collected from 19 longwall mines representing approximately 60% of all Australian longwall operations. In addition, six monitoring sites incorporated an array of hydraulic stress cells to measure the change in vertical stress throughout the various phases of the longwall extraction cycle. The sites also incorporated extensometers to monitor roof and rib performance in response to the retreating longwall face.
The study found strong relationships between the tailgate stability factor, the CMRR, and the installed level of primary support. The final outcome of the project was a chain pillar design methodology called Analysis of Longwall Tailgate Serviceability (ALTS). With ALTS, the recommended chain pillar width (rib to rib) is contingent upon an appropriate level of primary support (Figure C3). That level of primary support in turn depends on (1) the orientation of longwall retreat in relation to the magnitude and direction of the major horizontal stress and (2) the use of standing secondary support along the length of the gate road.
References:
Colwell MG, Frith R, Mark C. Analysis of Longwall Tailgate Serviceability (ALTS): A chain pillar design methodology for Australian conditions. In: Mark C., Heasley K, Iannacchione A, Tuchman R, eds. Proceedings of the Second International Workshop on Coal Pillar Mechanics and Design. NIOSH IC 9448, 1999, pp. 33-48.
Mark, C. Pillar Design Methods for Longwall Mining. BuMines IC 9247, 1990, 53 pp.
Mark, C. and A. T. Iannacchione. Analysis of Longwall Pillar Stability (ALPS)-An Update. Paper in Proceedings, Workshop On Coal Pillar Design and Mechanics, BuMines IC 9315, 1992, pp. 238-250.
Mark, C., F. E. Chase (1994). Design of Longwall Gate Entry Systems Using Roof Classification. Paper in New Technology for Longwall Ground Control: Proceedings of the USBM Technology Transfer Seminar, USBM SP 94-01, pp. 5-18.
Preventing pillar squeezes, massive pillar collapses, and bumps is critical to the safe pillar recovery operations. To help prevent these problems, the National Institute for Occupational Safety and Health (NIOSH) developed the Analysis of Retreat Mining Pillar Stability (ARMPS) computer program.
The goal of ARMPS is to help ensure that the pillars developed for future extraction (production pillars) are of adequate size for all anticipated loading conditions. The key is to be able to estimate the magnitudes of the various loads that they might experience throughout the mining process. The formulas used in ARMPS are based on those originally developed for the Analysis of Longwall Pillar Stability (ALPS) method which is widely used for longwall pillar design (see Appendix C). In ARMPS, the formulas were extensively modified for the variety of mining geometries typically found in pillar recovery operations.
The ARMPS program can model the significant features of most retreat mining layouts, including angled crosscuts, varied spacings between entries, barrier pillars between the active section and old (side) gobs, and slab cuts in the barriers on retreat (Figure D1). Other parameters that must be defined include the depth of cover, the mining height, the width of the entries, and the crosscut spacing.
ARMPS calculates stability factors (SF) based on estimates of the loads applied to, and the load-bearing capacities of, pillars during retreat mining. Four loading conditions are available, depending on the extent and location of nearby gob areas. The strength of an individual pillar (SP) is determined using a new pillar strength formula (the Mark-Bieniawski formula) that considers the effect of pillar length. The new pillar strength formula was needed because the strength of rectangular pillars can be significantly greater than square pillars, due to the greater confinement generated within them.
A database of 140 pillar retreat case histories has been collected across the United States to verify the program. It was found that satisfactory conditions were very rare when the ARMPS SF was less than 0.75. Conversely, very few unsatisfactory designs were found where the ARMPS SF was greater than 1.5. Analyses also indicate that that the ARMPS SF may be less meaningful when the depth of cover exceeds 750 ft (Figure D2).
References
Mark, C. and F. Chase (1997). Analysis of Retreat Mining Pillar Stability (ARMPS). Proc. New Technology for Ground Control in Retreat Mining, NIOSH IC 9446, pp. 17-34.
Mark, C. (1999). Empirical Methods for Coal Pillar Design. Proceedings of the Second International Workshop on Coal Pillar Mechanics and Design. NIOSH IC 9448, 1999, pp. 145-154.
Rating Sheets and Misc
1. CMRR field data sheet - Prince - Main Deep near 15W - page 1 of 1
2. Unit Rating (UR) Calculation Sheet - Prince - Main Deeps W 15 W - Unit No. 2 - silty sh
3. Unit Rating (UR) Calculation Sheet - Prince - Main Deeps W 15 W - Unit No. 1 - UC
4. Roof Rating (CMRR) Calculation Sheet - Prince - Main Deeps near 15W
5. Lithology Log of Seam - Smoky River - No. 4 Seam - 5B Mine
6. CMRR Calculation from Core - Prince
7. Roof Rating (CMRR) Calculation Sheet - Prince - Core (1)
8. Roof Rating (CMRR) Calculation Sheet - Prince - Core (2)
9. CMRR field data sheet - Smoky River - 9A Mine - page 1 of 1
10. Unit Rating (UR) Calculation Sheet - Smoky River - 9A - Highwall - Unit No. 1 - MST
11. Unit Rating (UR) Calculation Sheet - Smoky River - 9A - Highwall - Unit No. 2 - SLTST
12. Roof Rating (CMRR) Calculation Sheet - Smoky River - 9-A - Highwall - Fracture Zones
13. Roof Rating (CMRR) Calculation Sheet - Smoky River - 9-A - Highwall - Normal Roof
14. CMRR Calculation from Core - Smoky River Coal
15. CMRR field data sheet - Quinsam - 2 North - Mouth of 9 Panel - page 1 of 1
16. Unit Rating (UR) Calculation Sheet - Quinsam 2 North - Overcast with Mouth of 9 Panel
17. CMRR field data sheet - Quinsam - 4 South - Fault-bottom of Mains - page 1 of 1
18. Unit Rating (UR) Calculation Sheet - Quinsam 4 South - Fault at bottom of mains
19. CMRR Calculation from Core - Quinsam 4 South
20. CMRR Calculation from Core - Quinsam 2 North - Core QM9 P3
21. CMRR Calculation from Core - Quinsam 2 North - Core QM2
22. Core Logging and Testing - Quinsam 2 North - Core QM2
23. Core Logging and Testing - Quinsam 2 North - Core QM9P3
24. Roof Rating (CMRR) Calculation Sheet - Quinsam 2 North - Core QM 2
25. Roof Rating (CMRR) Calculation Sheet - Quinsam 2 North - Core QM9 P3
26. Roof Rating (CMRR) Calculation Sheet - Quinsam 2 North - Core QM 1
27. Memo from P. Cain (Smoky River Coal) re: core logging results and other things
28. Roof Borehole Logging Sheet - Quinsam - Borehole No. QM2 - 2 North - 4 Mains
CMRR field data sheet - Prince - Main Deep near 15W - page 1 of 1
UNIT RATING (UR)
CALCULATION SHEET
Mine Name Prince Date
Location Main Deeps W 15 W Data Collected by CM
1) Calculate the Individual Discontinuity Rating Unit No. 2 - silty sh
Discontinuity
Set 1 Set 2 Set 3
Cohesion-Roughness (table 1) 16
+ + +
Spacing-Persistence (table 2) 9
-----------------------------------
Individual Discontinuity Ratings 25
2) Enter the lowest of the Individual 25
Discontinuity Ratings
+
3) If there is more than one Discontinuity 0
set, enter the Multiple Discontinuity
Adjustment from table 3. Otherwise, enter 0. +
4) Calculate the Unit Strength (table 4) 15
+
5) Calculate the Unit Moisture (table 5) -3
(this applies only to Unit 1, or if upper
Unit is exposed to water)
-----
37 = Unit Rating (UR)
UNIT RATING (UR)
CALCULATION SHEET
Mine Name Prince Date
Location Main Deeps W 15 W Data Collected by CM
1) Calculate the Individual Discontinuity Rating Unit No. 1 - UC
Discontinuity
BDG Rooting
Set 1 Set 2 Set 3
Cohesion-Roughness (table 1) 16 10
+ + +
Spacing-Persistence (table 2) 9 17
-----------------------------------
Individual Discontinuity Ratings 25 27
2) Enter the lowest of the Individual 25
Discontinuity Ratings
+
3) If there is more than one Discontinuity -5
set, enter the Multiple Discontinuity
Adjustment from table 3. Otherwise, enter 0. +
4) Calculate the Unit Strength (table 4) 10
+
5) Calculate the Unit Moisture (table 5) -7
(this applies only to Unit 1, or if upper
Unit is exposed to water)
-----
23 = Unit Rating (UR)
ROOF RATING (CMRR)
CALCULATION SHEET
Mine Name Prince Date
Location Main Deeps near 15W Data Collected by CM
1) Calculate the UR Unit Thickness
weighted average (ft)
of the Unit Ratings
(RRw) 1. 23 x 2.5 = 57.5
+
2. 37 x 2.5 = 92.5
+
3. 23 x 1.0 = 23.0
+
4. x =
----------------
6 173
Bolted Interval (BI) ------ = 29 (RRw)
(m (in)) 6 (BI)
29 (RRw)
2) Calculate Strong Bed Difference (SBD)
Largest (UR) = 37 Strong Bed (SB)
37 (SB) - 29 RRW = 8 (SBD)
+
3) Calculate the Strong Bed Adjustment 2
(table 6)
+
4) Calculate the Unit Contact Adjustment 0
(table 7)
+
5) Calculate the Ground Water Adjustment 0
(table 8)
+
6) Calculate the Surcharge Adjustment 0
(table 9)
----
31 = CMRR
Lithology Log of Seam - Smoky River - No. 4 Seam - 5B Mine
CMRR CALCULATION FROM CORE
Mine: Prince
Unit: Underclay
Value Rating
Discontinuity RQD<25 20
Diametral - -
Axial PLT - 10*
Moisture Sensitivity -7
UNIT RATING ..................... 23
Unit: Sandstone
Value Rating
Discontinuity SPC=2.5-24" 33-39*
Diametral - -
Axial PLT -
UCS 7,000 psi 15
UNIT RATING ..................... 50
* Value assumed.
ROOF RATING (CMRR)
CALCULATION SHEET
Mine Name Prince Date
Location Core Data Collected by CM
1) Calculate the UR Unit Thickness
weighted average (ft)
of the Unit Ratings
(RRw) 1. 23 x 6 = 138
+
2. 50 x 2 = 100
+
3. x =
+
4. x =
----------------
8 238
Bolted Interval (BI) ------ = 30 (RRw)
(m (in)) 8 (BI)
30 (RRw)
2) Calculate Strong Bed Difference (SBD)
Largest (UR) = 50 Strong Bed (SB)
50 (SB) - 30 RRW = 20 (SBD)
+
3) Calculate the Strong Bed Adjustment 5 = (18 * 0.3)
(table 6)
+
4) Calculate the Unit Contact Adjustment -2
(table 7)
+
5) Calculate the Ground Water Adjustment 0
(table 8)
+
6) Calculate the Surcharge Adjustment 0
(table 9)
----
33 = CMRR
Assumes 8' Bolts
ROOF RATING (CMRR)
CALCULATION SHEET
Mine Name Prince Date
Location Core Data Collected by CM
1) Calculate the UR Unit Thickness
weighted average (ft)
of the Unit Ratings
(RRw) 1. 23 x 5 = 115
+
2. 50 x 3 = 150
+
3. x =
+
4. x =
----------------
8 265
Bolted Interval (BI) ------ = 33 (RRw)
(m (in)) 8 (BI)
30 (RRw)
2) Calculate Strong Bed Difference (SBD)
Largest (UR) = 50 Strong Bed (SB)
50 (SB) - 33 RRW = 17 (SBD)
+
3) Calculate the Strong Bed Adjustment 8 = (13 * 0.6)
(table 6)
+
4) Calculate the Unit Contact Adjustment -2
(table 7)
+
5) Calculate the Ground Water Adjustment 0
(table 8)
+
6) Calculate the Surcharge Adjustment 0
(table 9)
----
39 = CMRR
Assumes 8' Bolts
CMRR field data sheet - Smoky River - 9A Mine - page 1 of 1
UNIT RATING (UR)
CALCULATION SHEET
Mine Name Smoky River - 9A Date
Location Highwall Data Collected by CM
1) Calculate the Individual Discontinuity Rating Unit No. 1 - MST
Discontinuity
BDG Jts
Set 1 Set 2 Set 3
Cohesion-Roughness (table 1) 16 16
+ + +
Spacing-Persistence (table 2) 9 25 *
-----------------------------------
Individual Discontinuity Ratings 25 41
2) Enter the lowest of the Individual 25
Discontinuity Ratings
+
3) If there is more than one Discontinuity -3
set, enter the Multiple Discontinuity
Adjustment from table 3. Otherwise, enter 0. +
4) Calculate the Unit Strength (table 4) 15
+
5) Calculate the Unit Moisture (table 5) 0
(this applies only to Unit 1, or if upper
Unit is exposed to water)
-----
37 = Unit Rating (UR)
= 35 in fracture zones
* Reduces to 13 in fracture zones.
UNIT RATING (UR)
CALCULATION SHEET
Mine Name Smoky River - 9A Date
Location Highwall Data Collected by CM
1) Calculate the Individual Discontinuity Rating Unit No. 1 - SLTST
Discontinuity
Bdg-Maj Bdg-Min Jts
Set 1 Set 2 Set 3
Cohesion-Roughness (table 1) 16 30 16
+ + +
Spacing-Persistence (table 2) 23 15 25 *
-----------------------------------
Individual Discontinuity Ratings 39 45 41
Fracture Zones
2) Enter the lowest of the Individual 39 29
Discontinuity Ratings
+
3) If there is more than one Discontinuity -4 -4
set, enter the Multiple Discontinuity
Adjustment from table 3. Otherwise, enter 0. +
4) Calculate the Unit Strength (table 4) 30 30
+
5) Calculate the Unit Moisture (table 5) 0 0
(this applies only to Unit 1, or if upper
Unit is exposed to water)
-----
65 = Unit Rating (UR)
= 55 in fracture zones
* In fracture zones, reduces to 13.
ROOF RATING (CMRR)
CALCULATION SHEET
Mine Name Smoky River 9-A Date
Location Highwall - Fracture Zones Data Collected by
1) Calculate the UR Unit Thickness
weighted average (m (in))
of the Unit Ratings
(RRw) 1. 36 x 0.8 = 28.8
+
2. 55 x 0.4 = 22
+
3. x =
+
4. x =
----------------
1.2 50.8
Bolted Interval (BI) ------ = 42 (RRw)
(m (in)) 1.2 (BI)
42 (RRw)
2) Calculate Strong Bed Difference (SBD)
Largest (UR) = Strong Bed (SB)
55 (SB) - 42 RRW = 13 (SBD)
+
3) Calculate the Strong Bed Adjustment 8
(table 6)
+
4) Calculate the Unit Contact Adjustment -2
(table 7)
+
5) Calculate the Ground Water Adjustment 0
(table 8)
+
6) Calculate the Surcharge Adjustment 0
(table 9)
----
48 = CMRR
ROOF RATING (CMRR)
CALCULATION SHEET
Mine Name Smoky River 9-A Date
Location Highwall - Normal Roof Data Collected by CM
1) Calculate the UR Unit Thickness
weighted average (m (in))
of the Unit Ratings
(RRw) 1. 37 x 0.8 = 29.6
+
2. 65 x 0.4 = 26
+
3. x =
+
4. x =
----------------
1.2 55.6
Bolted Interval (BI) ------ = 46 (RRw)
(m (in)) 1.2 (BI)
42 (RRw)
2) Calculate Strong Bed Difference (SBD)
Largest (UR) = 65 Strong Bed (SB)
65 (SB) - 46 RRW = 19 (SBD)
+
3) Calculate the Strong Bed Adjustment 13
(table 6)
+
4) Calculate the Unit Contact Adjustment -2
(table 7)
+
5) Calculate the Ground Water Adjustment 0
(table 8)
+
6) Calculate the Surcharge Adjustment 0
(table 9)
----
57 = CMRR
CMRR CALCULATION FROM CORE
Smoky River Coal
Core tested by Peter Cain
Mudstone
Value Rating
Discontinuity No Data -
Diametral 780 psi 25
Axial PLT 4,120 psi 12
UCS Test 5,120 psi
UNIT RATING ..................... 37
Siltstone
Value Rating
Discontinuity No Data -
Diametral PLT 2,210 psi 42
Axial PLT 12,720 psi 22
UCS Test 11,300 psi
UNIT RATING ..................... 64
CMRR field data sheet - Quinsam - 2 North - Mouth of 9 Panel - page 1 of 1
UNIT RATING (UR)
CALCULATION SHEET
Mine Name Quinsam 2 North Date 9 Nov. 99
Location Overcast with Mouth of 9 Panel Data Collected by CM
1) Calculate the Individual Discontinuity Rating Unit No.
Discontinuity
Set 1 Set 2 Set 3
Cohesion-Roughness (table 1) 27 16
+ + +
Spacing-Persistence (table 2) 9 24
-----------------------------------
Individual Discontinuity Ratings 36 40
2) Enter the lowest of the Individual 36
Discontinuity Ratings
+
3) If there is more than one Discontinuity 0 -4
set, enter the Multiple Discontinuity If fractures
Adjustment from table 3. Otherwise, enter 0. + present
4) Calculate the Unit Strength (table 4) 18
+
5) Calculate the Unit Moisture (table 5) 0
(this applies only to Unit 1, or if upper
Unit is exposed to water)
-----
54 = Unit Rating (UR)
= CMRR
Note: Because the roof was just one unit, and no groundwater was observed,
UR = CMRR.
CMRR field data sheet - Quinsam - 4 South - Fault-bottom of Mains - page 1 of 1
UNIT RATING (UR)
CALCULATION SHEET
Mine Name Quinsam 4 South Date 9 Nov. 99
Location Fault at bottom of mains Data Collected by CM
1) Calculate the Individual Discontinuity Rating Unit No.
Discontinuity
Set 1 Set 2 Set 3
Cohesion-Roughness (table 1) 30
+ + +
Spacing-Persistence (table 2) 26
-----------------------------------
Individual Discontinuity Ratings 56
2) Enter the lowest of the Individual 56
Discontinuity Ratings
+
3) If there is more than one Discontinuity 0
set, enter the Multiple Discontinuity
Adjustment from table 3. Otherwise, enter 0. +
4) Calculate the Unit Strength (table 4) 22
+
5) Calculate the Unit Moisture (table 5) 0
(this applies only to Unit 1, or if upper
Unit is exposed to water)
-----
78 = Unit Rating (UR)
= CMRR
CMRR CALCULATION FROM CORE
Mine: Quinsam 4 South
Core: QM1
Unit: Sandstone
Value Rating
Discontinuity No Data -
Diametral 4,250 psi 53
Axial PLT/UCS 9,030 psi 15
UNIT RATING ..................... 68
CORE LOGGING AND TESTING
Core: QM1 - No log available.
Tested by CANMET
Unit: Sandstone - No RDQ or fracture SPC data
Axial PLT = 55.9 MPa = 8,110 psi
UCS Tests = 68.6 MPa = 9,950 psi
Diametral PLT = 29.3 MPa = 4,250 psi
CMRR CALCULATION FROM CORE
Mine: Quinsam 2 North
Core: QM9 P3
Unit: Lower Mst
Value Rating
Discontinuity RQD=60,2.4" 29
Diametral PLT 870 psi 25
Axial PLT 2,640 psi 12
UNIT RATING ..................... 35
Unit: Upper Mst
Value Rating
Discontinuity SPC=7.8" 36
Diametral PLT ~1,650 psi 33
Axial PLT 6,370 psi 15
UNIT RATING ..................... 48
CMRR CALCULATION FROM CORE
Mine: Quinsam 2 North
Core: QM2
Unit: Lower Mst
Value Rating
Discontinuity RQD=~50 29
Diametral 835 psi 25
Axial 6,060 psi 15
UNIT RATING ..................... 40
Unit: Upper Mst
Value Rating
Discontinuity RQD=~90 36
Diametral 1,770 psi 33
Axial/UCS ~8,800 psi 15
UNIT RATING ..................... 48
CORE LOGGING AND TESTING
Mine: Quinsam 2 North
Core: QM2 - Logged by S. Forgeron
Tested by CANMET
Lower Mst - 0-0.86 m (No fractures noted on log): RQD=52
Axial PLT testing 41.8 MPa = 6,060 psi
Diametral PLT 5.8 MPa = 835 psi
Upper Mst - 0.86-2.87 m: RQD=88
Axial PLT = 64.4 MPa = 9,340 psi
UCS testing = 57.1 MPa = 8,280 psi
Diametral PLT = 12.2 MPa = 1,770 psi
CORE LOGGING AND TESTING
Mine: Quinsam 2 North
Core: QM9P3 - Logged by P. Cain
Tested by P. Cain
Lower Mst - 0-0.66 m: RQD = 60
Fracture SPC = 6 cm = 2.4"
Axial PLT 18.2 MPa = 2,640 psi
Diametral PLT 6.0 MPa = 870 psi
Upper Mst: RQD = 94
Fracture SPC = 19.5 cm = 7.8"
Axial PLT = 43.9 MPa = 6,370 psi
Diametral PLT = 11.4 MPa = 1,650 psi
ROOF RATING (CMRR)
CALCULATION SHEET
Mine Name Quinsam 2 North Date
Location Core QM 2 Data Collected by
1) Calculate the UR Unit Thickness
weighted average (m (in))
of the Unit Ratings
(RRw) 1. 40 x .86 = 34.4
+
2. 48 x 1.6 = 76.8
+
3. x =
+
4. x =
----------------
2.5 111.2
Bolted Interval (BI) ------ = 44 (RRw)
(m (in)) 2.5 (BI)
44 (RRw)
2) Calculate Strong Bed Difference (SBD)
Largest (UR) = 48 Strong Bed (SB)
48 (SB) - 44 RRW = 4 (SBD)
+
3) Calculate the Strong Bed Adjustment 3
(table 6)
+
4) Calculate the Unit Contact Adjustment 0
(table 7)
+
5) Calculate the Ground Water Adjustment 0
(table 8)
+
6) Calculate the Surcharge Adjustment 0
(table 9)
----
47 = CMRR
ROOF RATING (CMRR)
CALCULATION SHEET
Mine Name Quinsam 2 North Date
Location Core QM9 P3 Data Collected by
1) Calculate the UR Unit Thickness
weighted average (m (in))
of the Unit Ratings
(RRw) 1. 35 x .66 = 23.1
+
2. 48 x 1.8 = 86.4
+
3. x =
+
4. x =
----------------
2.5 109.5
Bolted Interval (BI) ------ = 44 (RRw)
(m (in)) 2.5 (BI)
44 (RRw)
2) Calculate Strong Bed Difference (SBD)
Largest (UR) = 48 Strong Bed (SB)
48 (SB) - 44 RRW = 4 (SBD)
+
3) Calculate the Strong Bed Adjustment 3
(table 6)
+
4) Calculate the Unit Contact Adjustment 0
(table 7)
+
5) Calculate the Ground Water Adjustment 0
(table 8)
+
6) Calculate the Surcharge Adjustment 0
(table 9)
----
47 = CMRR
ROOF RATING (CMRR)
CALCULATION SHEET
Mine Name Quinsam 4 South Mine Date
Location Core QM 1 Data Collected by
1) Calculate the UR Unit Thickness
weighted average (m (in))
of the Unit Ratings
(RRw) 1. x =
+
2. x =
+
3. x =
+
4. x =
----------------
Bolted Interval (BI) ------ = (RRw)
(m (in)) (BI)
68 (RRw)
2) Calculate Strong Bed Difference (SBD)
Largest (UR) = Strong Bed (SB)
(SB) - RRW = (SBD)
+
3) Calculate the Strong Bed Adjustment 0
(table 6)
+
4) Calculate the Unit Contact Adjustment 0
(table 7)
+
5) Calculate the Ground Water Adjustment 0
(table 8)
+
6) Calculate the Surcharge Adjustment 0
(table 9)
----
68 = CMRR
Memo from P. Cain (Smoky River Coal) re: core logging results and other things
Roof Borehole Logging Sheet - Quinsam - Borehole No. QM2 - 2 North - 4 Mains