Datasets:

Raneechu

/

unsuper3

like 0

rocesses such as swelling or shrinking of magma chambers tectonic activities which may alter the stress fields on a very large scale thereby causing an increase or shift in the direction of geostatic stresses processes that created the slope these may include creep on the slope or creep in weak strata below the foot of the slope factors that contribute to low or reduced shear strength include the following factors stemming from the initial state or inherent characteristics of the material these factors include the presence and orientation of discontinuities slope orientation the existence of massive beds over weak or plastic materials and the alternation of permeable beds and weak impermeable beds changes in shear strength due to weathering and other physiochemical reactions these changes can include softening of fissured clays physical disintegration of granular rocks due to the action of frost or by thermal expansion and contraction hydration or dehydration of clay materials drying of clays which results in cracks drying of shales which creates cracks on bedding and shear planes and removal of cement within discontinuities by solution changes in intergranular forces due to water content and pressure in pores and fractures which may result from rapid drawdown of a lake or reservoir rapid changes in the elevation of the water table rise of the water table in a distant aquifer or seepage from an artificial source of water the primary affect of groundwater pressure in reducing the stability of rock slopes is the resulting decrease in effective shear strength of discontinuities changes in structure can be caused by remolding clays or clay like materials upon disturbance by the fissuring of shales and preconsolidated clays and by the fracturing and loosening of rock slopes due to the release of vertical or lateral restraints upon excavation miscellaneous causes can include weakening of a slope due to progressive creep or due to the actions of roots and burrowing animals data collection collecting adequate and appropriate data for stability analysis is a key aspect of slope design obtaining incorrect results from slope stability analysis is predominantly the result of failing to analyze the critical failure mode or not having the suitable estimates of the input parameters such as rock strength or the geometry of geologic structures with the use of computers our ability to construct mathematical models and perform the calculations exceeds our ability to collect adequate input data for the models there are two aspects to the problem of data collection sampling and measurement as an example consider the specific task of determining the uniaxial compressive strength of the rock mass this is usually measured by conducting a series of compressive strength tests on cylinders 50 75 mm 2 3 in in diameter and 100 200 mm 4 8 in in length cored from the rock mass these tests should be conducted ac

cording to astm international standards astm 2008 the population of interest referred to as the target population by statisticians includes all the cylinders of the given size contained within the potential slope instability mass because of stress considerations the volume may very well extend some distance beyond the projected limits of the slope instability boundary it is obvious that all of the target population could not be tested so the strength distribution must be estimated using the test results from some small hopefully representative sample population the availability of samples for testing is determined by access which would be the ground surface the pit wall underground workings and drill holes where there is no preexisting pit or underground workings and the ground surface is covered by alluvium access to samples is restricted to drill holes these accessible samples are referred to as the sampled population the samples that are actually collected and tested are referred to as the sample population to make valid statistical inferences about a population every member of the population in question must have an equal likelihood of being sampled and the tested samples must be an unbiased representation of the population an example of sampling bias is the determination of rock quality designation rqd from rock cores from vertically drilled boreholes that do not intersect an important vertical joint set the rqd will be biased due to underrepresentation of that important joint set in the case of geologic structure data collection parameters such as orientation length and spacing are geometric rather than scalar and cannot be measured at a point this results in a window problem particularly in the case of fracture length if the fracture is larger than the observation window such as a bench face the length cannot be directly measured this is why surface mapping is preferable to drillhole data where the core diameter is the window there is also an orientation bias as a linear sampling window such as a drill hole does not intersect fractures parallel to the window data collection should be well organized with specific objectives regarding the use of the data and the quantity required collecting data for data s sake should be avoided as it will result not only in information that is not used but the possible omission of information needed ongoing data reduction is important in order to determine whether a sufficient quantity of appropriate data is being collected the following general procedure is suggested for geologic data collection prior to or during a slope stability investigation kliche 1999 1 determine the boundaries between geologic materials with different properties weathering can be used as criteria e g color changes hammer impact look for differences in sedimentation e g grain size distribution consider differences in joint intensity e g high versus less 2 determi

ne structural features folds faults and faulting systems bedding in sedimentary rocks schistosity and cleavage in metamorphic rocks discontinuities e g joints shear zones bedding planes faults 3 map discontinuities figure 8 3 17 shows a typical discontinuity survey data sheet location coordinates along a line survey or x y z values determine material type i e host rock note the type of discontinuity fault shear zone bedding plane etc determine the orientation of the discontinuity dip and dip direction or dip and strike determine the persistence or the continuity of the discontinuity i e look at the continuity of the discontinuity and the joint length determine joint intensity number of joints per unit distance normal to the strike of the set and the degree of separation between the joint faces determine the openness or closed nature of the discontinuity if it is open look for the presence of filling material such as gouge or transported material if filling is present look at the type clay granular material crystalline material or veins measure the thickness of the filling material determine the waviness and the roughness i e first and second order of asperities 4 conduct a sampling program obtain samples of intact rock both weathered and unjointed obtain samples of the discontinuity surface for shear testing care must be taken so as to not disturb the two surfaces of the discontinuity if filling material is present obtain representative samples may require soil mechanics tests for strength parameters 5 determine groundwater conditions locate springs or seeps determine the permeability of the rock and the joints joint conductivity is usually the most important may need drawdown or recovery tests measure the discharge of springs or seeps at least twice per year geology and major structure conventional geology provides the distribution of rock types and alteration and the location of major structures geologic data should be in the form of a surface map cross sections and level maps it is preferable to have two sets of documents the factual sheets that show only the actual observations and a set of interpreted maps and cross sections for the design of final pits a geologic map of a trial pit design and cross sections normal to the pit wall should be constructed rock fabric rock fabric is the orientation length and spacing of fractures these are the geometric attributes used in stability analysis and in characterizing the rock mass on a pit scale the number of fractures such as joints are too numerous to map fracture mapping therefore consists of measuring the attributes of a subset of the total fractures and characterizing the population with distributions of the attributes detailed mapping reveals that the orientation of fracture sets dip and dip direction or dip and strike has a normal or bivariate normal distribution since the orientation is vecto

r quantity it is properly a spherical normal distribution this can be seen by observing the contoured pole density clusters of figure 8 3 18 the measurable aspect of joint size is the trace length which is the intersection of the joint and the mapping surface the negative exponential appears to be the best distribution for trace lengths on the basis of fit to mapping data and theoretical considerations models such as the circular disk and the poisson flat have been postulated to describe joints in three dimensions these models can be used to correct for the observation window limitation several common mapping methods are available call 1992 fracture set mapping this is a modification of conventional joint mapping where fracture sets are identified by eye and the orientation length and spacing are recorded if joints or other structure orientations have been recorded during regular geologic mapping they can be compiled and used in slope design detail line the detail line method is a systematic spotsampling method in which a measuring tape is stretched along the bench face or outcrop to be measured for all the fractures along the tape the point of intersection with the tape orientation length roughness filling type and thickness are recorded figure 8 3 17 to get an adequate representation of the fabric at least 100 fractures should be mapped this is the least subjective method as individual fractures are recorded and it provides the most detailed length and spacing data it is relatively inefficient however as more observations are made on closely spaced fracture sets than are required for adequate statistical representation cell mapping in this method mapping surfaces such as a bench face are divided into cells normally the width of the cells is made equal to the height of the cells within each cell the fracture sets are identified by eye and the orientation length and spacing characteristics are recorded cell mapping is a combination of fracture set mapping and detail line with the efficiency of visual identification of fracture sets and some of the more rigorous measurements of detail line oriented core oriented core provides information on fracture orientation and spacing but the length of fractures cannot be directly measured this technique is used when the rock types of interest are not exposed it domains mapped on the surface extend back or downward for some distance nicholas and sims 2000 rock strength properties since the spatial variability of rock properties is large the potential for sampling error is greater than the measurement error for this reason it is preferable to use simple test methods for a number of samples than to use an expensive precise method on one sample for the shear strength of discontinuities and fault gouge the direct shear test is recommended as it is a simulator of field conditions since the shear normal failure curve may be nonlinear it is imp

ortant to use normal stress values for the test that represent the expected range of normal stresses for potential failure geometries in the slope the tests at each normal stress should be run with sufficient displacement to obtain both the peak shear strength and the residual shear strength as the residual shear strength usually is a better estimate of in situ strength than the peak strength for intact rock unconfined compression and brazilian disk tension tests are recommended in addition to obtaining the compressive and tensile strengths the uniaxial compression tests can be gauged with strain gauges to obtain the young s modulus and poisson s ratio for the intact rock index tests such as the point load can also be used to evaluate the spatial variability of intact rock strength for the rock mass where direct testing is not possible indirect methods such as the rock structure rating rock mass rating or the rock tunnel quality index and back analysis must be used for a more complete discussion of rock strength properties refer to chapter 8 4 hydrology the most important groundwater parameter for stability purposes is the groundwater pressure distribution within slopes this distribution can be obtained in two ways 1 by direct measurement of pressure using piezometers or 2 by determining pressures from an analysis of the hydraulic properties of the rock mass e g geology and permeability characteristics the most satisfactory approach is usually to measure groundwater pressures with piezometers at representative locations and to correlate these data with analytical studies based on a thorough understanding of the geology and on selected permeability or conductivity measurements of representative soil and rock strata in slope stabilization using water pressure control the focus is on reducing groundwater pressure typically any stabilization involves installing pressure control devices dewatering pumps drainage galleries or drain holes at a number of locations and reducing the water level i e the pressure in the material around and between those devices the process focuses on obtaining a major water pressure reduction in an acceptable time period standard hydrologic procedures such as piezometers and pump tests can be used to obtain the current pore pressure distribution and the permeability for predicting changes in pore pressure with time and changes in pit geometry simple techniques such as measuring the water level in drill holes can be used two factors need to be considered however 1 water behavior in rock slopes is a fracture flow phenomenon and porous media analysis while useful at a regional scale may be a poor predictor of pore pressure at pit slope scale and 2 the critical factor in slope design is the pore pressure rather than the quantity of water a low permeability rock mass may yield very little water and appear dry yet have significant pore pressure stress measurements a number

of techniques are available for measuring in situ stress generally each technique should be evaluated prior to its use at a particular mine site during early site investigations when no underground access is available the only practical method for measuring in situ stresses is by hydrofracturing haimson 1978 in which the hydraulic pressure required to open existing cracks is used to estimate in situ stress levels after underground access is available overcoring techniques for in situ stress measurement leeman and hayes 1966 worotnicki and walton 1976 can be used with such techniques as the commonwealth scientific and industrial research organization csiro type hollow inclusion cell and provided that sufficient care is taken in executing the measurements the results are usually adequate for design purposes design steps in slope design include 1 defining design sectors 2 conducting a bench design analysis to determine the maximum interramp slope 3 conducting interramp design analysis using economic criteria for the selection of interramp angles and 4 evaluating the resulting overall slope for potential instability and modifying the design if required slope design is an interactive process because a trial pit is required to select design sectors but the development of a trial pit requires slope angles frequently the way this is handled is to generate a trial pit shell using critical parameters based on experience bench height bench angle berm width interramp angle overall pit angle as more geotechnical geological hydrological and operating information comes in the trial pit is refined over and over until the final pit design is established rarely however is there a final pit design until the end of mining as the design parameters will change as additional or more current information comes in to the designer design sectors to conduct stability analyses and develop optimum slope angles for input into pit design the proposed pit must be divided into design sectors that are sections of the pit with similar geologic and operational characteristics figure 8 3 19 the first criterion for the selection of design sectors is the structural domain which is an area within which the rock properties and fabric are consistent typical structural domain boundaries are lithologic contacts and major structures that separate areas of dissimilar fabric the second criterion is wall orientation since rock is usually anisotropic different wall orientations within the same structural domain can have significantly different modes of instability and different optimum angles one way of handling the situation of a change in highwall orientations with the rock properties and fabric remaining constant is to divide the sector into subsectors based on the changing orientation a third criterion for defining design sectors is operational considerations sectors containing critical structures such as in pit crushers conveyor

s and haul roads require different stability criteria than the same wall orientation in the same structural domain without the critical structures that is the design factor of safety may have to be modified upward based on exposure and risk considerations for each of the design sectors or subsectors the rock fabric major structure orientation data and planned highwall orientation can be plotted on a stereographic projection this diagram is used to determine which failure modes may be kinematically possible and to select structure sets for stability analyses catch bench berm design bench faces are normally mined as steeply as possible so that some bench scale rockfalls and raveling can be expected development of catch benches in mine slopes is necessary in areas of such rockfall since the catch benches if designed properly prohibit rocks from rolling from upper portions of the pit slope to the working areas where personnel and equipment are located the bench geometry defines the steepest interramp slope that can be mined while maintaining adequate catch bench widths the two primary factors that control bench configuration are the type of mining equipment that is used and the bench face angles that can be achieved the type of mining equipment determines the safe operating height of the bench the achievable bench face angles are controlled by rock strength geologic structure characteristics and the mining techniques used to construct the slope e g the blasting and digging practices ryan and pryor 2000 bench design is the process of conducting stability analyses to estimate the bench face angles selecting the bench width and to a limited extent the bench height the bench height is controlled by the height of the mining levels but it is possible to increase the height by leaving catch benches on every other level double benching or every third level triple benching the sole purpose of a catch bench is to catch rocks so that they do not continue unhindered to the toe of the slope or away from the base of the rock cut the bench is designed with a specific width relative to its height so that rocks will come to rest before falling off the next crest in addition a back break distance is often incorporated into the design as the bench crest will often fail from its usual vertical position the design catch bench width is thus always wider than the bench width required for safety based on an analysis of rockfall mechanics ritchie 1963 developed width and depth criteria for a ditch at the toe of a slope to protect highways from rockfalls falling rocks impact close to the toe of the slope but because of horizontal momentum and spin can roll considerable distances from the toe the concept of ritchie s design was that the rock would impact in the ditch and the side of the ditch would stop the horizontal roll a reliability approach to catch bench design wherein the analysis is structured to evaluate the pe

rcentage of the slope area that meets or exceeds a chosen catch bench width criteria has been recently developed ryan and prior 2000 the developers have found this technique to be a more useful risk management approach for rockfall containment and for slope management in this approach a combination of structural modeling bench face stability analysis and the modified ritchie criteria is used to determine the catch bench reliability which refers to the percentage of benches having final widths equal to or greater than the modified ritchie criteria the developers state that the selection of the proper reliability for maintaining a catch bench of a certain width is dependent on such factors as the potential for slope raveling the proximity to large slope failures the decision to contain overbank from a higher pushback on the benches the length of time the benches are expected to be functional the climate the type of blast control and the operator s experience because the bench configuration is based on the reliability of the catch bench width it is the amount of local back break rather than the bench face angle that is of most concern the percent reliability therefore represents the percentage of the bench along a given level that would be wider than the minimum required bench width to catch rockfalls the reliability should be selected on the basis of the potential for rockfall and the exposure of personnel and equipment for example the catch bench in raveling ground above a haul road requires a greater reliability than catch benches in a stripping area with more competent ground in practice reliabilities from 60 90 have been satisfactory interramp slope angle for a given bench height and corresponding bench width the upper limit of the interramp angle becomes a function of the bench face angle the bench face angle however is not a unique value as variability of the rock fabric results in varying amounts of back break back break is defined as the distance from the design crest to the as mined crest figure 8 3 23 because of this variability it is preferable to use a reliability approach rather than using the mean bench face angle calculating an interramp slope using the minimum bench width and the mean bench face angle results in 50 of the benches being too narrow the procedure is to select a percentage reliability and use the cumulative frequency distribution of the bench face angle to find the angle where the percentage greater is equal to the reliability figure 8 3 24 this gives the design bench face angle to use with the minimum bench width and the bench height to calculate the interramp slope angle in an operating property the actual bench faces can be measured and the measured bench face angle distribution can be used in design where existing bench faces are not available a bench face angle distribution can be obtained by running a stability analysis for a vertical face for thi

s analysis the plane shear wedge and step path analyses are run using the fracture data the height analysis should be incremented in steps up to the bench height and the resulting back break composited as short fractures that would not result in full bench failure can still cause crest back break this bench face angle distribution is referred to as the theoretical bench face distribution as the effect of blasting and digging is not included if there is a strong geologic control such as bedding or foliation the measured and theoretical bench face angles are the same where no strong structure exists the theoretical bench face angles should be reduced to include the effect of blasting based on comparisons that have been made between measured and theoretical angles the reduction should be between 10 and 20 depending on the controlled blasting to be used interramp design the stability of interramp slopes is primarily controlled by intermediate and major structure failure geometry where major structures can be specifically located in space the geometry relative to the slope can be defined and a discrete stability analysis can be conducted commonly however the number of mapped structures is large and the distance between the mapping sites and the design wall is greater than the length of the structures in this case the structural data must be considered a statistical representation of the structures that will occur in the design slope and a probabilistic analysis is required to obtain the input for stability analysis the wall orientation can be plotted on a lower hemisphere stereographic plot of the poles of the fractures and the major structures the fractures and major structures are sorted into design sets based on their orientation relative to the orientations for failure modes and the distribution of orientation length and spacing can be computed for the design set these design sets may not correspond to geologic sectors although the boundaries of the sets may be adjusted to avoid splitting a geologic sector an advantage of this approach is that it is based on kinematic tests for viable failure geometry making it unnecessary to test all the structures for each failure mode major structures in the case of throughgoing major structures where the geometry is known a safety factor can be calculated for specific slope angles and slope heights using analytical models described in the references for the appropriate failure model for a deterministic design the slope angle with the desired factor of safety would be selected in the reliability method the probability of sliding can be calculated by monte carlo sampling of the shear strength distribution to obtain a distribution of safety factors and computing the area of the safety factor distribution 1 other techniques can be used such as the point estimate method harr 1984 or calculating the probability that the shear strength is less than the strength required

for a safety factor of 1 because of the variability of the shear strength a safety factor 1 is used to reduce the risk of instability to an acceptable level one problem with this is that a given safety factor will have a different level of risk depending on the dispersion of the input parameters the advantage of the reliability approach is that it deals directly with the risk failure volume estimation where the geologic structures compose a statistical population the probability of failure for the single occurrence of a specified failure mode is a function of the probability that the structures exist and form a viable failure geometry as well as the probability of sliding marek and savely 1978 the probability of existence is calculated from the orientation length and spacing of the structures to calculate the expected number of failures and the expected failure volume for input to a cost benefit analysis the probability of failure for the possible failure modes must be calculated for a range of heights and angles and then composited cost of failure given the expected number of failures and the expected failure tonnage the cost of slope failure can be estimated failure costs consist of cleaning up failure material repairing haul roads repair of facilities lost production due to disruption of operations the value of lost ore buried by a failure and engineering costs the method used to estimate failure cost is a what if mine planning procedure a failure is postulated for a design sector a plan for responding to the failure is made and the cost of the plan is estimated these exercises are useful whether or not a full cost benefit optimization is done as they can lead to modifications of the mine plan that will reduce the impact of slope instability overall slope the overall slope is usually flatter than the interramp slope because of catch benches or other step outs thus the overall slope normally will be more stable than the interramp except for stress induced failure or failure modes not analyzed for the interramp numerical modeling techniques can be used in addition to deterministic techniques to analyze the potential for the various modes of failure including block flow multibench wedge planar type and rotational numerical techniques have been time consuming and expensive in the past however with the faster computers and better software currently available it has become very feasible and routine a quick check can be made for block flow potential using the charts developed by coates 1981 if the charts do not indicate block flow potential with any regional stress assumption numerical modeling is not needed unless there is a high contrast in stiffness between adjacent materials in the slope the charts assume a homogeneous material and would therefore not indicate stress concentrations produced by stiffness contrasts changes in the overall slope angle have relatively little effect on the stress conce

ntration at the toe of the slope where a greater concentration could produce block flow therefore block flow potential would not be a suitable method for selecting overall slope angles a more effective design approach would be to design the slope based on other criteria and to make provision in the mine plan for step outs if needed in the toe area of the pit to reduce the stress concentration produced by the notch effect of the bottom of the slope the loss of ore from step outs at the toe would have less economic impact than the amount of stripping required to have the same effect on block flow potential rotational shear analysis should be run for the overall slope even on rock slopes to verify that it would not be a critical failure mode rotational shear would be a primary method of analysis for both interramp and overall slopes in alluvium and low rock mass strength slopes such as soft coal measures the general surface analysis should be used for the overall slope to evaluate mixed mode failure types where part of the failure is structurally controlled and part is failure of low rock mass strength nondaylighted wedge and plane shear failures in which the weak rock at the toe fails are becoming recognized as a more significant failure mode this is in part because pits are becoming steeper and deeper and partly because more pits have been designed for the simpler sliding block failure modes slope support and stabilization slope stabilization techniques can be divided into six general categories 1 grading 2 controlled blasting 3 mechanical stabilization 4 structural stabilization 5 vegetative stabilization 6 water control grading grading involves the shaping of the rock slope into a more stable configuration it may include flattening the slope leaving benches in the slope face or sculpting the slope face to a more natural appearance catch bench design and application are as discussed previously controlled blasting specialized blasting techniques are used to control overbreak and produce a competent final excavation wall by minimizing damage to the final pit walls from production blasts the purpose of overbreak control is to achieve a stable highwall by limiting the damage from production blasting beyond the cut limit often a secondary purpose is to achieve an aesthetically appealing wall several drilling and blasting techniques have been developed for overbreak control to accomplish this purpose mckown 1984 floyd 1998 including the following blasting the energy level is decreased adjacent to the wall to reduce overbreak this decrease in energy level is often achieved for competent rock simply by reducing the charge weight in the row nearest the slope by about 30 60 floyd 1998 for less competent rock masses additional modifications to the blast design may be required to minimize overbreak damage these modifications can include using decked charges reducing the burden i e the distance to the nearest

free face and spacing of the last row minimizing subdrilling and increasing the delay interval between the last two rows of blastholes the primary advantage of the modified production blasting technique is that it requires few design changes the primary disadvantage is that the wall rock is not protected from crack dilation gas penetration and block heaving floyd 1998 reducing the number of holes per delay will reduce the peak particle velocity but for the perimeter row of holes and the buffer row a production hole charge is usually too large and must be reduced to maintain the same powder factor the hole spacing must be reduced concurrently with the reduction in hole charge in practice this method of controlled blasting increases the measured bench face angle by 5 savely 1986 presplit blasting presplitting uses lightly loaded closely spaced drill holes that are fired before the production blast to form a fracture plane across which the radial cracking from the production blast cannot travel konya 1995 as a secondary benefit the fracture plane formed may be aesthetically appealing figure 8 3 25 shows a mine highwall where presplit blasting has been utilized to control overbreak and produce a stable final wall the presplit blast may be fired a considerable amount of time before the production blast hours days weeks or longer or shortly before as on a prior delay delayed blasting techniques are frequently used to separate the detonation times of explosive charges i e individual holes or series of holes as in a row of holes trim cushion blasting trim blasting is a control blasting technique that is used to clean up a final wall after the production blast has taken place konya 1995 the trim blast may be on a later delay of the production blast or at a much later date possibly years after blasting the purpose of the trim blast is twofold to create an aesthetically appealing final wall and to enhance the stability of the final wall by removing overbreak from the production blasting because the trim row is shot after the final production row the trim blast does little to protect the stability of the final wall from production blasting it does however provide enhanced stability by removing the loose material caused by overbreak from the production blasting in trim blasting cost is incurred from extra drilling longer blasthole loading time and some reduced mine production but these short term costs are often offset by decreased future costs in terms of stripping and slope failure both of these potential future costs stripping and slope failure can be incurred as a result of overbreak into the final highwall from the production blasting line drilling line drilling is often included as one of the controlled blasting techniques however it is not a blasting technique per se line drilling uses a single row of unloaded closely spaced drill holes at the perimeter of the excavation when a produc

tion blast adjacent to a series of line drilled holes is detonated the shock wave from the detonating holes will cause a stress concentration around the unloaded linedrilled holes if the stress exceeds the rock strength failure will occur in the form of a crack extending from one linedrilled hole to the adjacent line drilled hole line drilling is an expensive perimeter control technique because of the number of holes required for the technique to work properly two aspects of final wall design should be considered in selecting the appropriate controlled blasting technique or combination of techniques for a particular job defining rock damage criteria and developing a procedure to design blasts that will minimize rock damage without seriously affecting production mechanical stabilization mechanical methods of slope stabilization are those that alter or protect the slope face to reduce erosion prevent rockfall or to reduce raveling common methods include protective blankets geotextiles and wire net or mesh protective blankets protective blankets made from jute excelsior burlap cotton or other natural or manufactured materials have been used for many years for erosion control and to prevent or reduce raveling on cut slopes the blankets are usually pinned to the slopes and combined with seed and fertilizer the purpose of pinning the blanket is to hold it in place until the vegetation takes root the blankets are often expected to deteriorate and thus biodegrade over time as the vegetation takes hold geotextiles a geotextile is defined by astm as any permeable textile material used with foundation soil rock earth or any other geotechnical engineering related material as an integral part of a man made project structure or system christopher and holtz 1985 geotextile applications can be divided into four primary functions separation drainage reinforcement and filtration christopher and holtz 1985 in separation layers of different sizes of solid particles are separated from one another by the geotextile e g landfill covers in drainage the geotextile allows water to pass in the special case of drainage transmission the geotextile itself acts as a drain to transmit water through soils of low permeability e g horizontal drains below heap leach pads in the case of reinforcement the geotextile acts as a reinforcing element in the earth through either stress distribution or an increase in soil modulus e g a net against rockfalls for filtration the fabric acts in a similar fashion to a two dimensional sand filter allowing water to move from the soil while retaining the soil e g silt screens wire net or mesh another method of slope stabilization involves draping or pinning wire netting over the slope face to prevent rockfalls from bouncing outward from the toe region three types of wire mesh are commonly used for this purpose 1 welded wire fabric such as that used in concrete reinforc

ement or 2 chain link mesh as is commonly used for fencing and 3 flexible wire rope nets with or without underlaid wire netting geobrugg type netting a typical welded wire mesh application would be to use mesh with a 100 100 mm 4 4 in or 150 150 mm 6 6 in opening and a wire size from 9 to 4 gauge seegmiller 1982 chain link fence is often coated with a galvanizing agent and will therefore better withstand adverse environmental conditions also because of the nature of its construction chain link fence tends to be more flexible and stronger pinning the net or wire mesh to the face holds the rock in place and reduces rock removal at the toe see figure 8 3 26 the pins typically rock bolts rock dowels thread bars or rebar must be strong enough and spaced close enough to hold large loose rocks and prevent them from dislodging and tearing the mesh structural stabilization structural stabilization includes those methods that reinforce the structure of the rock at the slope face or provide a structure that supports the slope methods available include the use of gunite or shotcrete rock bolting and construction of rock buttresses or retaining walls shotcrete one common method of mechanical stabilization uses pneumatically applied mortar and concrete generally known as gunite or shotcrete sprayed or pumped onto the slope face to seal the face and bind together small fragments on the face this approach is used primarily to prevent weathering and spalling of a rock surface as well as to knit together the surface of a slope generally for rock slope stabilization the material is applied in one 50 75 mm 2 3 in layer brawner 1994 one disadvantage of shotcrete is its low tensile strength for this reason welded wire mesh anchored to the rock is often used to reinforce the shotcrete a problem with using wire mesh as reinforcement for shotcrete is the difficulty of molding the mesh to a rough surface where the surface is irregular large gaps may develop between the mesh and the rock making bonding of the shotcrete to the rock difficult additives can be added to either the wet or dry mix to provide additional strength and durability steel fibers when added to the mix increase the tensile strength of the shotcrete by providing numerous bonding surfaces within a small area the fiber reinforcement also reduces the risk that shrinkage cracks will develop during curing in many cases the addition of fibers can replace wire mesh as reinforcement thus reducing the overall cost steel reinforcement members steel reinforcement in the form of rock bolts cable bolts resin grouted thread bars or rock dowels are used to tie together the rock mass so that the stability of a rock cut or slope is maintained rock bolts are commonly used to reinforce the surface or near surface rock of the excavation and rock anchors are used for supporting deep seated instability modes in which sliding or separation on a di

scontinuity is possible a rock anchor generally consists of a bar or cable of highstrength steel tensioned inside a borehole to about 60 70 of its yield strength tension in the member is transmitted to the surrounding rock mass by anchorage points at the ends the length of the rock anchor can be from 3 m 10 ft to 100 m 330 ft sage 1977 resin grouted tensioned thread bar and grouted cables provide a means to control large failure blocks lengths of the units may be as short as 3 m 10 ft or as long as 100 m 330 ft depending on the specific application holes for installation of the anchors are normally drilled well past the potential failure plane then the anchors are inserted and grouted in place with or without tensioning grade 60 423 7 mpa or 60 000 psi resin grouted tensioned thread bar comes in diameters ranging from 19 1 mm 0 750 in to 57 3 mm 2 26 in and in nominal lengths of 12 2 m 40 ft such thread bar can be cut to lengths shorter than 12 2 m 40 ft or made longer by coupling two or more units together grouted cables were introduced to mining for reinforcement of the backs of cut and fill stopes cable bolting whether tensioned or untensioned is widely used in mining applications the cable bolts should be made from high strength steel about 1 380 mpa or 200 000 psi yield strength typically because the steel will creep in tension as a result a gradual decrease in anchor load will occur over time this loss in strength is approximately the same for all types of steel sage 1977 a special prestressing jack figure 8 3 27 is required for tensioning the bolt also for the tensioned cable bolt a special bearing plate an anchor block and cable gripping cones are required for a more complete discussion of steel reinforcement members that may be used in slope reinforcement refer to chapter 8 8 vegetative stabilization vegetative techniques are most frequently used for aesthetic purposes such as slope reclamation however there are many treatment methods that use vegetation to improve the stability of a slope generally these methods are most successful when minor or shallow instability such as raveling or erosion is involved as is usually the case for soil slopes or highly fractured rock slopes buss et al 1995 the establishment of vegetation on steep soil slopes or loose rock slopes is often enhanced by the construction of benches or stair step terraces in the slope face these arrangements act to hold the seed mix in place to encourage infiltration and to impede water flow in order to minimize erosion and sedimentation natural or manufactured mats webs or fabrics can also be used for erosion control and to hold seeds in place though their cost and effectiveness often limit their use they require high labor inputs for installation and cost much more than tacked or hydromulched straw in addition some are not well adapted to fitting to rough surfaces they must also be heavy en

ough or anchored in enough spots to prevent wind whipping the planting of trees or other large woody plants on rock slopes is beneficial in softening the appearance of the cut slope to make it appear more natural buss et al 1995 in order to make access easier for planting trees or shrubs benches berms or furrows must be constructed in the slope face if access allows the use of a tree spade enables the transplanting of large mature trees the plants are placed in holes that have been previously excavated on the reclamation site a single hole drilled and blasted in rock will provide an excellent site for a transported tree or shrub the plants are transplanted with a minimum of root disturbance tree spades however are expensive to use and their use should be reserved to transporting hard to establish trees or to achieve an objective of immediate stocking with mature trees water control the presence of groundwater in a rock slope can have a detrimental effect on stability for the following reasons water pressure reduces the stability of the slopes by diminishing the shear strength of potential failure surfaces water pressure in tension cracks reduces stability by increasing the driving forces changes in moisture content of some rock particularly shales can cause accelerated weathering and decreases in shear strength freezing of groundwater can cause wedging in waterfilled fissures due to temperature dependent volume changes in the ice this can result in expansion of an opening such as a tension crack upon freezing which does not contract upon thawing of the water erosion of weathered rock by surface water and of lowstrength infillings by groundwater can result in local instability where the toe of the slope is undermined or a block of rock is loosened excavation costs can be increased when working below the water table by far the most important effect of groundwater in a rock mass is the reduction in stability resulting from the water pressures within the discontinuities grading and shaping are major considerations in the control of surface water surface water can be controlled through a combination of topographic shaping and runoff control structures glover et al 1978 surface water allowed to flow down a slope or to pond on benches of a slope can infiltrate into the ground along discontinuities and thereby cause an increase in the driving forces on an unstable area through a buildup in pore pressure runoff control structures include dikes waterways diversion ditches diversion swales and chutes or flumes glover et al 1978 the purpose of these structures is to intercept surface water flow before it reaches a critical area and to divert it to a disposal area the purpose of subsurface drainage i e groundwater control is to lower the water table and therefore the water pressure to a level below that of the potential failure surfaces methods of subsurface drainage include drain holes pu

mped wells and drainage galleries or adits drain holes horizontal drain holes drilled into the face of the slope from the toe region offer an effective method of slope drainage normally the holes are 50 150 mm 2 6 in in diameter and are drilled at an inclination of 3 to 5 from the horizontal the length of the holes should extend beyond the critical failure surface the direction of the drain holes depends on the orientation of the critical discontinuities the optimum design is to intersect the maximum number of significant discontinuities for each unit length spacing of the drain holes can range from about 7 to 30 m 20 to 100 ft and lengths into the slope should not exceed one half the slope height with a minimum length of 15 m 49 ft and a maximum length of 100 125 m 330 410 ft brawner 1982 for high rock cuts installation of drain holes at different levels is recommended where rock is taken out in several lifts drain holes should be drilled at the toe of every lift brawner 1994 dewatering wells dewatering wells are designed primarily to lower the groundwater level to a predetermined depth and to maintain that depth until all below ground activities have been completed the main purposes for construction dewatering include the following driscoll 1986 intercepting seepage that would enter an excavation site and interfere with construction activities improving the stability of slopes thus preventing sloughing or slope failures preventing the bottoms of excavations from heaving because of excessive hydrostatic pressure improving the compaction characteristics of soils in the bottoms of excavations drying up borrow pits so that excavated materials can be properly compacted in embankments reducing earth pressures on temporary supports and sheeting drainage galleries or adits drainage adits or galleries driven under a pit or into a slope or highwall to intercept the groundwater can provide an effective method of drainage where employed drain holes should be drilled from the adit upward in a fan pattern to increase drainage effectiveness an adit can be used not only for drainage but also as a means of obtaining detailed discontinuity information this type of data is useful for future slope stability predictions additional information such as water quality and the variations in permeability along the length of the adit can also be gathered the cost of driving an adit is high however it need only be large enough to allow efficient excavation and to properly drain the problem area generally 1 0 1 5 2 0 2 2 m 3 3 4 9 6 5 7 2 ft slope management three general principles of slope mechanics should be kept in mind in cases of slope instability call and savely 1990 1 slope failures do not occur spontaneously one or more of the forces acting on a potentially unstable rock mass must change in order for the mass to become unstable 2 most slope failures tend toward equilibrium a slope fails

because it is unstable under the existing conditions failure tends to bring the slope to some sort of equilibrium it normally involves a reduction in the driving forces and or an increase in the resisting forces of the failed zone 3 a slope failure does not occur without warning prior to failure measurable movement and or the development of tension cracks will occur these indications of failure can develop indicating imminent slope failure then subside for a long period of time indicating apparent stability in the evaluation of the necessity and type of stabilization technique to be used the first issue to be considered is the degree of urgency if the slope has started to move immediate remedial actions should be taken these actions may include evacuating structures closing roadways or cordoning off a section of a mine after the situation has become less urgent determining the cause of the instability is necessary by the following means visual observation based on experience water level measurements slope instrumentation tests on the materials and a survey of the discontinuity patterns golder 1971 laboratory tests may be required on the rock and discontinuities to determine appropriate strength parameters detection and monitoring of instability when a rock or soil mass is disturbed either by the actions of people or by natural events it undergoes a redistribution of stresses resulting in a change in shape this readjustment is reflected in displacements deflections pressures loads stresses and strains which can be detected and measured many of the same measurement methods and instrumentation techniques can also be used to investigate the mechanical properties of the mass the interaction between the mass and any associated artificial structures and the effectiveness of remedial measures proposed to correct defects in either the mass or the structures monitoring is the surveillance of engineering structures either visually or with the aid of instruments brown 1993 the objectives of a rock slope monitoring program are as follows call 1982 to maintain a safe operation for the protection of personnel and equipment to provide advance notice of instability thus allowing for the modification of the excavation plan to minimize the impact of the instability to provide geotechnical information in order to analyze the slope failure mechanism design appropriate remedial measures and or conduct a redesign of the rock slope monitoring can be done from within the rock mass and or on the excavation boundary the techniques available to measure the various components of rock deformation may be placed in two general categories observational and instrumentation windsor 1993 observational techniques include simple visual observations photographic recording and electronic and optical surveying instrumentation techniques include the application of mechanical and electronic instruments such as extensometers

inclinometers strain gauges and crack gauges an overview of both observational and instrumental techniques as well as the types of instruments for monitoring deformation is given in table 8 3 1 an effective slope monitoring program consists of the systematic detection measurement interpretation and reporting of evidences of slope instability measurements are normally made of both surface and subsurface displacement in order to provide an accurate assessment of slope instability surface displacement surface displacement measurement by means of observational techniques surveying by conventional methods automatic surveillance and or global positioning system and or instrumentation techniques movement indicators extensometers inclinometers convergence indicators is the preferred monitoring method a combination of the methods and instruments should be used as no one method would give the entire picture tension crack mapping tension cracks are an early obvious indication of instability by systematically mapping the cracks the geometry of a failure can be better defined all cracks should be mapped regardless of apparent cause often cracks that appear to be the result of local bench failure or blasting form a pattern showing an impending larger failure when plotted on a pit map the ends of the cracks should be flagged or marked so that on subsequent visits new cracks or extensions of existing cracks can be identified wireline extensometers portable automated wireline extensometers can be used to provide monitoring in areas of active instability across tension cracks these monitors can be quickly positioned and easily moved the extensometer should be positioned on stable ground behind the last visible tension crack and the wire should extend out to the unstable area for warning devices or for information on deformation within a sliding mass wire extensometers can be placed at any strategic location anyone working in the area can make an immediate check on slope movement by inspecting the instruments the automated slope monitoring system has the ability to record data automatically and then transmit the data to a central computer system for analysis the system is used for real time displays of slope movement as well as for long term analysis of all recorded information martin 1996 the automated system is composed of two main components the slope monitor unit which is located in the field and the central computer and radio which may be located in the mine office the system may be programmed to transmit alarms if certain conditions occur such as broken wire excessive movement or velocity or communication failures for instance if the wire breaks or the slope anchor probe pulls loose the wire spool falls to the base of the mounting tripod the falling spool pulls a magnet off the electronics box which immediately radios a warning to the central computer martin 1996 the central computer system collect

s and processes the data and generates screen displays and or reports on slope movement status a solar panel can be attached and used to recharge the battery power supply survey monitoring the most widely used method of monitoring for movement employs the theodolite edm electronic distance meter total station or programmable robotic theodolite in conjunction with an array of monitoring prism targets this continues to provide the most detailed movement history in terms of displacement directions and rates in the unstable areas locations of the targets should be chosen so that relative movement of the unstable area can be monitored additionally the permanent control point s from which the targets are shot i e observed must be located on stable ground outside the slide area and within view of the targets the network should be a set of well conditioned triangles with each vertex point being visible from two other points and the length of each line of sight within the measuring range of the equipment windsor 1993 the instrument as well as any backsights used should be located on stable ground away from the slide area s computerized automatic theodolites are gaining popularity to monitor movement at 100 or more survey prism stations the robotic theodolite may be housed in a small shed or dispatch building at a vantage point in direct line of sight with all prism stations a computer system may also be housed in the robotic slope monitoring shed this system uses a specialized software package to control the theodolite which after preprogramming shoots the array of prisms automatically on a programmed cycle after completion of a cycle the system may be set to shoot one or more subcycles or may start another cycle of shooting the set of prisms readings can be trans ferred via hard wire internet modem or radio telemetry link to a base computer located elsewhere for further processing slope stability radar the slope stability radar figure 8 3 28 is a state of the art technology used for slopestability monitoring of open cut mine walls it provides continuous precise and real time on line measurement of rock wall movements across the entire face of a wall it remotely scans a rock slope to continuously monitor the spatial deformation of the face using differential radar interferometry the system is used to detect deformation movements of a rough wall with submillimeter accuracy and with high spatial and temporal resolution the effects of atmospheric variations and spurious signals can be reduced via signal processing means the advantage of the slope stability radar over other monitoring techniques is that it provides full area coverage without the need for mounted reflectors or equipment on the wall in addition the radar waves adequately penetrate through rain dust and smoke to give reliable measurements in real time and 24 hours a day groundprobe 2009 subsurface displacement surface displacement measure

ments do not determine the subsurface extent of instability although it is possible to make inferences from displacement vectors there are many situations where measurement of subsurface displacement is needed these measurements are commonly made using borehole inclinometers or borehole extensometers borehole inclinometers an inclinometer measures the change in inclination or tilt of a borehole and thus allows the distribution of lateral movements to be determined versus depth below the collar of the borehole as a function of time wilson and mikkelsen 1978 therefore the application of inclinometers to slope stability studies is important for the following reasons to locate shear zone s to determine whether the shear along the zone s is planar or rotational to measure the movement along the shear zone s and determine whether the movement is constant accelerating or decelerating two types of inclinometer are in common use the in place inclinometer and the traversing probe type of inclinometer the in place inclinometer was developed for automated monitoring it is composed of a string of inclinometer sensors permanently mounted in the casing the sensors are normally positioned within the casing to span the zone where movement is anticipated the string of sensors is usually attached to a data acquisition system that is programmed to trigger an alarm if certain boundary conditions are exceeded the inplace inclinometer is expensive its use is generally limited to only the most critical applications boisen and monroe 1993 the traversing probe type of inclinometer was developed to address the problem of expense it employs a single sensor that can be used to monitor any number of inclinometer casings tilt readings are typically obtained at 2 m intervals 2 ft intervals are used with probes that use the u s customary system of measure as the probe is drawn from the bottom to the top of the casing the main drawback to the system is that it is slow and requires an on site operator boisen and monroe 1993 borehole extensometers the fixed borehole extensometer measures only axial displacement between a fixed number of reference points on the same measurement axis when more than two reference points are used the instruments are referred to as multiple position or multipoint extensometers multipoint extensometer data can reveal the relative movement between anchor points and the distribution of displacement in addition to the magnitude rate and acceleration of displacement slope indicator company 1994 the basic components of a fixed borehole extensometer are an anchor a linkage and a reference head the reference head is installed at the borehole collar the linkage system may be composed of wires or of solid rods it spans the distance between the reference head and the anchor a change in this distance indicates that ground movement has occurred measurements are taken at the reference head with a depth microm

eter or an electronic sensor and are used to determine the displacement precision reliability and cost the number of different devices that can be used for monitoring as well as the precision and sophistication of the devices are a function of the ingenuity time and budget of the engineer in charge of monitoring since none of these factors is infinite hard choices must be made some general guidelines for decision making follow measure the obvious things first surface displacement is the most direct and most critical aspect of slope instability tension cracks can easily be mapped or photographed the surveyor can set survey points and log movement on a regular basis simpler is better the reliability of a series system is the product of the reliability of the individual components a complex electronic or mechanical device with a telemetered output to a computer has significantly less chance of being in operation when needed than do two stakes and a tape measure precision costs money the cost of a measuring device is often a power function of the level of precision measuring to 10 mm 0 4 in is inexpensive compared to measuring to 0 001 mm 0 0004 in redundancy is required no single device or technique tells the complete story a single extensometer or survey point will not indicate the area involved in the instability and if it is destroyed the continuity of the record is lost multiple survey prisms should be placed in and around a slide area that is being monitored using optical or robotic surveying techniques timely reporting is essential data collection and analysis must be rapid enough to provide information in time to make decisions modern computerized data acquisition systems are not always available therefore persons doing the monitoring and reporting should submit promptly any indication of movement acceleration requires attention establish a monitoring schedule a definite monitoring schedule should be established the frequency of monitoring is a function of the precision of the system the rate of movement and how critical the area is if there is heavy rain or a large blast in the area additional measurements should be made cooperation between operations and engineering is important equipment operators often have an intuitive feel for ground conditions any changes in the condition of an area observed by operators e g an observation of the development of tension cracks should be reported to engineering for follow up data reduction and reporting should be conducted the following measurements or calculations should be made for each survey reading date of reading incremental days between readings and total number of days the survey point has been established coordinates and elevation magnitude and direction of horizontal displacement magnitude and plunge of vertical displacement magnitude bearing and plunge of resultant total displacements velocity and any change in velocity

acceleration deceleration of the movement vector both incremental and cumulative displacement values should be determined calculating the cumulative displacement from initial values rather than from summing incremental displacements minimizes the effects of occasional survey aberrations slope displacements are best understood and analyzed when the monitoring data are graphically displayed for engineering purposes the most useful plots are horizontal position vertical position elevation versus change in horizontal position plotted on a section oriented in the mean direction of horizontal displacement displacement vectors cumulative total displacement versus time and incremental total displacement rate velocity usually in feet or meters per day versus time all graphics should be kept up to date and should be easily reproducible for ease of distribution by studying several graphs simultaneously the movement history of a particular slope can be determined precipitation data should also be recorded in order to evaluate possible correlations with slope displacement a rain gauge or system of rain gauges located at the mine site can be used to measure occurrences and amounts of precipitation in addition measurement of the average daily temperatures will provide some indication of freeze and thaw periods the location of mining areas and the number of tons mined should also be recorded on a regular basis because slope displacements are often associated with specific mining activity blasting records including seismic records should be kept and referenced to the unstable areas a histogram can be made of tons mined versus time and this plot can then be compared to the total displacement graphs a formal monthly slope stability report should be prepared containing at least the following data maps plan view showing the unstable areas and locations of monitoring devices displacement vectors blast locations tension crack map piezometric surface plots and graphs cumulative displacement versus time displacement rate versus time precipitation versus time mining versus time prediction of time to failure for critical unstable areas an action plan for each critical unstable area time dependent slope movement characteristics all excavations whether natural or human made deform with time in response to excavation according to zavodni 2000 the most commonly observed evidence of time dependent deformation of cut slopes is the development of tension cracks behind the slope crest formation of cracks on the slope and toe heave zavodni further states that mining operations can proceed safely with minimum interruption if failure mechanisms are understood and slopes are properly monitored a serious slope instability condition is usually accompanied by gradual development of one or more tension cracks behind the slope crest this situation then normally allows for time displacement monitoring surface displac

ement measurements employing prism targets are usually adequate for monitoring slope movement the prisms may be accompanied by extensometers closely monitored to determine slope movement velocities in order to predict slope behavior recent advances in survey equipment robotic theodolites and real time transmission of data allow increasingly precise monitoring of movement accompanied by immediate data display and analysis regressive and progressive movement if it is to occur a conventional open pit slope failure begins after the initial response of the excavation and is normally associated with the creation of one or more tension crack s at or near the crest of the slope the development of such cracks is evidence that the slope is at limiting equilibrium and the driving forces stress just equals or exceeds the resisting forces stress in the case of the slope at limiting equilibrium stability is decreased by increasing the driving forces stresses decreasing the resisting forces stresses or by changing both the driving and resisting forces stresses whether the failure is regressive or progressive depends on whether a potential or active rock slope failure tends to become more stable or less stable a regressive failure is one that shows short term decelerating displacement cycles if disturbing events external to the rock are removed from the slope environment a progressive failure on the other hand is one that will displace at an accelerating rate usually an algebraically predictable rate to the point of collapse unless active and effective control measures are taken zavodni and broadbent 1982 a third type defined by zavodni and broadbent is what is known as the regressive progressive condition or transitional system these curves are shown on figure 8 3 29 type i regressive condition the regressive failure type is shown as curve a on figure 8 3 29 this failure type is characterized by a series of either accelerating or decelerating displacement trends as revealed from continuous monitoring programs the characteristic that qualifies this curve as regressive is the deceleration of each cycle between external stimuli the cycles are believed to be initiated when the driving force stress temporarily exceeds the resisting force stress thereby causing the rock slope condition to drop slightly below a safety factor of 1 the velocity of movement will decay if the external disturbance is eliminated the excess driving force is usually related to an external event such as a mine blast earthquake precipitation event temperature change groundwater pressure change or excavation of buttressing rock characteristics of regressive type failures are when the ratio of driving stress to resisting stress decreases with displacement and slope will tend to become more stable with time and show decelerating or stick slip behavior type ii progressive condition the progressive failure demonstrates an increase in th

e rate of displacement over time until collapse as shown by curve b in figure 8 3 29 decelerating cycles may be present but would be subtle and nearly indistinguishable from the long term trend the time period over which progressive displacement of a large scale failure takes place is usually short 4 to 45 days zavodni 2000 type iii regressive progressive condition as shown by curve c in figure 8 3 29 a regressive type failure may transition into a progressive type failure and rapidly lead to collapse causes of this change in behavior can include a situation where mining daylights a sliding surface breakup or excavation of rock at the toe of a slope or an increase in water pressure slide management some degree of slope instability can be expected with virtually any slope cut in rock whether the slope is a mine highwall is for a road cut or is a part of some other construction project when slope stability investigations indicate that the possibility of slope failure exists there are a number of response options available call and savely 1990 leave the unstable area alone continue mining without changing the mine plan unload the slide through additional stripping leave a step out i e a bench or berm of unexcavated rock at the toe of the slope to increase the resisting forces of the slope conduct a partial cleanup mine out the failure support the unstable ground dewater the unstable area the option or combination of options that is chosen depends primarily on the nature of the instability and its operational impact each situation should be evaluated separately with safety aspects carefully considered and with the cost of any remedial action and benefit included contingency planning mine planning should have the flexibility to respond to slope instability rather than an after the fact crisis response to forced deviation from a rigid mine plan contingency plans should be prepared in advance so that the response to slope instability is well thought out a management philosophy of do something even if it s wrong will frequently result in more problems than the original event i e slope instability precipitated also the lack of adequately trained personnel for the geotechnical group will result in significant scheduling problems and production delays when a slope failure event occurs operational flexibility should be built into the mining plan for example adequate ore should be uncovered and available so that production is not dependent on a single location there should be more than one access road into the pit for service and haulage vehicles the loss of the haul road when it is the only access into the mine or a portion of the mine will shut down production for an extended period of time whenever possible more than one access to working benches should be maintained and labor and equipment should be available for slide cleanup when necessary rock mechanics is the appl

ied science of the mechanical behavior of rock and rock masses in a stress field in mining knowledge of rock mechanics is combined with empirical reasoning to make predictions of performance of excavations so that safety environmental impact and economic value can be optimized this chapter focuses on the applied part of the definition and particularly how the science can be applied to the practicalities of mining it provides a ready reference for some of the key definitions and methods in current use the opportunity is taken to provide some guidelines on the use of various concepts highlighting where major uncertainties are still present the coverage of the subject is not detailed or complete but is presented to introduce the subject to nonspecialist mining engineers numerous publications provide the detail the theoretical basis and recent case studies to which the specialist geotechnical engineer is referred e g hoek et al 1998 brady and brown 2004 there continue to be major advances in both mining rock mechanics and mining engineering mining practices have evolved in the face of increasing expectations of workplace safety and environmental impact zero harm is now the stated goal of most mining companies greater reliance is now placed on bulk mining systems such as sublevel and block caving in metal mines and draglines and longwalls in the coal sector the costs of delays in bulk mining systems can be very high there are now greater demands that the operations perform rigorously to both plan and budget in this environment monitoring of mining systems is an essential part of ensuring workplace safety but it is not adequate to ensure continued production or appropriate environmental impact better forecasts of mining conditions are required and these demand a much better understanding and application of rock mechanics in the science of rock mechanics there have been major developments in computer based stress analysis such that twoand three dimensional 2 d and 3 d stress analyses of openings in rock masses are now routine the ease at which these can be conducted together with the associated graphics may give an unjustifiable level of confidence in the predicted performance of the excavations although knowledge of the stresses around a mine opening is essential it is not sufficient to ensure adequate performance one key objective of the mining process is the control of displacements into and around the openings failure of rock and rock masses is readily induced around mining openings so the prediction of displacements requires knowledge of the strength of rock masses unfortunately knowledge of rock strength is comparatively poor with fundamental revisions being made within the last decade much of rock mechanics depends on the appreciation of engineering geology it is only by understanding the distribution and characteristics of rock types the associated discontinuities and the size of the blocks def

ined by them that rock mechanics can be applied nature of rock rock and rock masses are inherently complex brady and brown 2004 the complexities relate to the following fractures are developed in a compressive stress field this limits the ability to transfer knowledge from other branches of engineering where failures are induced in a tensile stress regime mine openings can vary in scale from the order of meters to hundreds of meters over this scale there is a need to consider both intact rock and the rock mass with all its discontinuities as a result of the discontinuities the tensile strength of a rock mass must be considered to be zero in most cases mining takes place below the original water table water may have an impact in terms of both effective stress and in terms of reducing the strength of certain rocks weathering of the rock around mine openings can lead to chemical and or physical alteration of the mineral components and hence rock strength the importance of the concept of scale is illustrated in figure 8 4 1 at the scale of a roadway or a bench the behavior of the excavation may be controlled by the strength of the intact portions or by the presence of one or two joints or beds at a large scale for example the wall of a deep surface mine the behavior of the rock mass may be controlled by the overall joint and bedding structure should the excavation intersect a major fault zone then the behavior will be controlled by that single feature for analytical convenience rock masses are often considered to be continuous homogeneous isotropic and linear elastic chile according to the following definitions continuous intact rock with no breaks homogeneous all the same rock isotropic rock properties are the same in all directions linear elastic an increment or decrement in stress produces the same increment or decrement of strain there is no permanent deformation and no failure in reality rock masses are discontinuous inhomogeneous anisotropic and nonlinear elastic diane the assumption of chile continuum behavior introduces major risks many of the operational mining problems relate to the rock mass not behaving as a continuum for example in many applications ground support is best addressed through the consideration of blocky rock as is bench design in surface mines though there are notable exceptions rock mechanics and mine design rock mechanics design fits within an engineering design methodology bieniawski 1993 steps in the methodology emphasize the need to characterize the site geology as well as the strength of the materials there is a need to simplify the complex conditions into geotechnical models concept formulation that can be subsequently analyzed there are five pathways through the concept formulation and analysis steps four of which rely on aspects of the science of rock mechanics figure 8 4 2 precedent practice is the exception this is still an important pa

thway so long as there is confidence that the ground conditions are the same rock mass classification schemes seek to transfer knowledge gained from previous mining experiences to other sites the schemes allocate numerical values to features of the rock mass considered likely to influence overall behavior and then combine them into a single rating value although useful for translating experience within a single mine site or mining district caution is required in their use in the absence of analysis of the mechanics of the problem being considered brady and brown 2004 the schemes are useful for planning purposes but they are not necessarily useful for the prescription of rock support during actual construction palmstrom and broch 2006 the term behavior model is used here to mean the identification of failure or collapse modes separate from the calculation of stresses wedge planar and toppling failures in the benches of surface mines are examples the limit equilibrium methods characteristic of soils engineering are also examples of how the failure mechanism is identified prior to the analysis kaiser and kim 2008 argued that by concentrating on the behavior of the excavation boundary the significance of brittle tensile failure was identified and a new failure criterion was developed numerical codes may be used to assist in the calculation of stresses in the numerical approaches the failure zones are identified after the stresses are calculated therefore they depend on the failure criterion input to the model field stresses constitutive equations stress strain relationships and failure criteria are required and these require assumptions simplifications of equal or greater magnitude as the formulation of behavior models some continuum numerical codes e g phase 2 flac fast lagrangian analysis of continua determine stresses about openings assuming the rock mass behaves as a continuum in some codes discontinuities can also be modeled in distinct element codes e g udec universal distinct element code the rock mass is treated as quasi rigid blocks that interact through movement along the discontinuities codes for both 2 and 3 d analyses are now in widespread use which pathway to use depends on the stage of the mining project it is good practice to use at least two pathways to provide an internal check on the design figure 8 4 3 provides general guidelines for each stage there will be situations where different levels of geotechnical knowledge will require different strategies precedent practice is the strongest pathway for operations but this should be accompanied by a way of confirming that the same geotechnical regime is present historically the latter has been done through the rock mass classification pathway and the limitations of this were mentioned earlier the classification schemes were developed before the advent of personal computers that make stress analyses readily available since 19

90 the focus has been on the numerical code pathways and there is now a realization that knowledge of the constitutive equations and failure criteria is limited the behavior model pathway may be more appropriate as it demands good observation of how the rock actually behaves and this not only identifies immediate hazards but also identifies the need for new scientific knowledge figure 8 4 3 suggests that the rock mass classification pathway is particularly suitable at early stages of a mining project when there is little data available at the prefeasibility level numerical codes can be used to determine stress redistributions around openings and they allow for optimization of the overall layout while at the same time work is done to characterize the discontinuity fields as the mine moves to the operational phase the focus needs to move to the performance of the excavation boundaries roadway or bench and the behavior of relatively small blocks of rock may be best addressed by the behavior model pathway sophisticated analytical and prediction tools do not ensure adequate or appropriate predictions using concepts from soils engineering there is a need to be aware that for a given quantum of data there is an appropriate degree of sophistication in the analysis beyond which the accuracy of the prediction may decline lambe 1973 in referring to lambe it is not suggested that accurate class a predictions are essential or even possible in mining ventures although the better the class a predictions the lower the risk to capital the message is that the mining engineer must maintain an ongoing awareness of factors that contribute to unsuccessful performance and introduce this awareness into comprehensive risk management tools morgenstern 2000 scope the application of four of the pathways requires knowledge of rock mechanics either of rock mass classification systems or of the combination of the strengths of rock and rock masses as well as the stresses around excavations this chapter is a general introduction to the subject of rock mechanics so as to provide a broad understanding of the issues and the associated uncertainties it is anticipated that the scope will be of value to mine planners seeking to understand the constraints that need to be applied to mining engineers who are using the readily available stress analysis codes and require an understanding of the input parameters and to engineering geologists that are looking for an introduction to the engineering concepts specialist geotechnical engineers are referred to many of the standard texts for detailed treatment of the subject rock properties there are two fundamental sets of rock properties the discontinuities and the substance itself this section describes how rocks can be characterized at the laboratory and outcrop scale the rock properties in the following discussion are used to different degrees in the four analytical pathways discussed earlier so they

must be determined as accurately as possible discontinuities the distinctive feature of rock masses is the presence of discontinuities in this context discontinuities are defined as features of a rock mass that have zero or negligible tensile strength this terminology allows a differentiation from textural features within intact rock from a geological perspective discontinuities can include joints bedding partings faults veins shear zones cleavage and schistosity the term defect can also be used interchangeably with discontinuity a corollary of this definition is that the tensile strength of a rock mass should be assumed to be zero the international society for rock mechanics isrm provides the framework for the description of discontinuities spacing and persistence table 8 4 1 presents the isrm definitions of terms for spacing and persistence although this table may be satisfactory for igneous and metamorphic regimes it does not adequately address the range of conditions that can be encountered in sedimentary sequences such as found in coal mining bedding partings can have persistence values very much greater than 20 m and there are situations in some of the coal mass mining regimes longwalls where it is necessary to describe bedding parting spacings in the order of 30 to 50 m in most rocks it appears that discontinuity spacing tends to follow a negative exponential function although in some cases a lognormal distribution better applies brady and brown 2004 in relatively undisturbed sedimentary rocks a valid first order assumption is that the mean joint spacing is equal to the mean spacing of bedding partings ji et al 1998 the rock mass is composed of cubes though this simplification should be applied with caution the rock quality designation rqd is a simple measure of discontinuity spacing and rock mass quality the rqd is the percentage of core that reports in lengths greater than 100 mm the core lengths should be measured along the center line and only double or triple tube drilling should be used reduced discontinuity spacing and greater persistence should be anticipated near faults many ground control problems that are reported in the vicinity of faults are related more to the smaller block size defined by the closer spaced discontinuities than to higher stresses that are often presumed to be present in fact the deviatoric stresses near faults are likely to be lower than in the adjacent rock mass orientation the onset of planar wedge and toppling failures in surface mines and also around underground openings is related to the relative orientation of the excavation to the dominant discontinuities the determination of the dip and dip direction strike of the discontinuities is therefore a critical step in any mine design mapping of any exposures should be conducted and supplemented with information from drill holes modern borehole imaging tools are to be preferred over core orientation to

obtain orientation data from drilling programs in some situations knowledge of the regional geology can assist in forecasting orientations for example in gently deformed sedimentary regimes it may be possible to infer joint orientations from knowledge of the orientation of fold axes fookes et al 2000 surface conditions joint roughness aperture and fill discontinuities in rock are rarely perfectly smooth and planar the undulations and roughness of natural discontinuity surfaces have a significant influence on their shear strength properties particularly at the low stresses around the immediate excavation boundary barton 1973 one set of definitions for surface conditions is presented in table 8 4 2 along with corresponding joint roughness coefficients jrcs the adjacent rock walls of a discontinuity may be separated by infill such as clay calcite and fault gouge the aperture of a discontinuity is the perpendicular distance between the adjacent rock walls when the intervening space is filled with air or water in most cases the aperture will be small unless there has been recent movement along irregular surfaces dilatant stress conditions or erosion of fill materials hydraulic conductivity with laminar flow is proportional to the cube of the aperture shear strength if a rough irregular discontinuity is considered between two blocks of rock held together by a stress normal to the surface and exposed to a shear stress as the shear stress increases shear displacement will increase eventually a peak shear stress will be reached peak strength and continuing shear displacement will require a lesser shear force residual strength repeating this test for different normal loads allows for the construction of two sets of envelopes for the normal stress shear stress pairs shear testing is normally done using a direct shear test analogous to the left hand diagram in figure 8 4 4 with a restricted number of laboratory tests e g three to five the peak strength envelope is typically found to be bi linear the angle of the second linear segment is called the basic friction angle jb and the gradient of the initial linear segment is related to both the basic friction angle and the roughness of the discontinuity intact rock the strength of intact rock is determined in the laboratory by a series of tests on rock cores the cores are installed in a purpose designed pressure vessel and exposed to a number of different confining pressures an axial stress is applied and as it increases the rock compresses axially and it initially may contract slightly laterally later in the test the rock will dilate laterally the relationship between stress and strain is not linear over the full range of applied stress sometimes a yield point can be defined where the stress strain curve deviates significantly from being linear the peak strength is measured for each confining stress and this increases with the magnitude of the confine

ment figure 8 4 5 at the same time the shape of the post peak part of the curve may change from brittle curve a through strain softening curve b to plastic curve c and strain hardening curve d as the confinement increases depending on the nature of the rock material uniaxial testing strength laboratory unconfined or uniaxial compressive strength ucs is the key index to rock strength the ucs is the strength measured in the laboratory on samples with the length diameter ratio greater than 2 and typically on 50 to 65 mm diameter samples testing should be conducted on fresh core and at a moisture content as close as possible to the in situ value control of moisture content is important especially for the lower strength rocks e g 20 mpa as it is these that may be easily overstressed in some mining conditions testing conducted at lesser aspect ratios will result in higher indicated strengths and testing of larger samples will result in lower strengths field estimates table 8 4 4 have been found to be very useful for the lower strength rocks that can be readily overstressed around openings and yet may be difficult to sample and test in the laboratory when using this table one should be careful not to include core breaks along incipient discontinuities or textures the ucs is a fundamental parameter in most strength criteria and should be measured directly whenever possible the engineering literature suggests that the ucs can be determined from the point load strength index ucs 24 is50 value which is the point load strength corrected to a 50 mmdiameter equivalent unfortunately the constant has been found to vary massively between various rock types and the default value of 24 should be used with extreme care the testing standards require a large number of tests at least twice as many as the ucs a high degree of scatter in the results is common point load testing should not be relied on for rocks with ucs values of less than about 25 mpa deformation parameters in mining rock mechanics a common assumption is made that prior to failure rock can be considered to be linear elastic in elasticity theory there are 21 constants in the constitutive equations that relate deformation to stress with increasing material symmetry the number of elastic constants is reduced an orthotropic material has three material axes and nine independent elastic constants a gneiss or granite may be orthotropic many sedimentary rocks show pronounced differences in properties of samples parallel and perpendicular to the bedding and can be considered to be transversely isotropic five elastic constants are needed for the description of transversely isotropic rock only two independent elastic constants are needed for the description of isotropic rock that by definition lacks directional mechanical properties the prefailure shape of typical stress strain curves for rock is not perfectly linear figure 8 4 6 assuming linearel

astic and isotropic behavior the elastic deformation parameters of young s modulus and poisson s ratio can be calculated from laboratory data the young s modulus is calculated as the ratio of change in axial stress to the corresponding change in axial strain and the poisson s ratio is the ratio of the change in the radial strain to axial strain typically tangent values are determined at 50 of the peak strength and average values are determined to be between one third and two thirds of peak strength young s modulus and poisson s ratio are assumed not to vary with confining stress secant values to peak strength may also be quoted the interrelationships between other elastic parameters are g e 2 1 n l 2ng 1 2n where g shear modulus also known as the modulus of rigidity e young s modulus n poisson s ratio l lame s constant it has been determined empirically that the ratio of the young s modulus to ucs the modular ratio varies within relatively narrow limits depending on rock type table 8 4 3 the layering in sedimentary rocks introduces a degree of transverse isotropy implementation of transverse isotropy requires two deformation moduli and two poisson s ratios as well as an independent shear modulus there is very little guidance on the choice of values of the independent shear modulus for continuum analyses of layered rock masses the author has found that values in the order of 100 to 250 mpa may be necessary to account for both textural impacts and bedding partings confined strength in the last 20 years and especially since 2000 there have been major changes in the models for the confined strength of rock as of 2009 there is no single failure criterion for the full range of stresses and separate analyses are needed for the excavation boundary and the far field the envelope of confining stress axial stress points for peak strength is nonlinear the more sample points that are collected the more the nonlinearity of the relationship becomes apparent the three common approaches to fitting a relationship are straight line the mohr coulomb criterion that is often used in software codes power hoek brown criterion hoek and brown 1980 multilinear the brittle criterion kaiser et al 2000 the testing cells used in standard laboratory tests mean that one confining stress is applied the test is conventionally referred to as a triaxial test with the understanding that the magnitudes of two of the stress axes are identical a consequence of this is that the failure criteria refer to only two stresses 1 and 3 because 2 3 discussed later a true triaxial failure criterion is the subject of ongoing research rock masses some mining problems are at such a scale that a large volume of the rock mass must be considered there are two approaches to assess the performance of rock masses one approach is the classical rock mass classification route that bypasses the need for any analysis of modes

of behavior a more recent approach has been to assume that the rock mass is so impacted by discontinuities that it can be considered to be a continuum and to then use a rock mass classification to generate equivalent strength and deformation parameter rock mass classification systems seek to apply a single numerical value to describe the condition of rock masses based on a consideration of what are believed to be key parameters controlling behavior two of the most popular systems are the rock mass rating rmr of bieniawski 1973 1976 1989 which has been revised several times and the q system barton et al 1974 in both systems the ucs of the rock plays a relatively minor part in the characterization reflecting the origin of the systems in what have been referred to as hard rock commentaries on the application of both systems often state that they are suitable for hard rock unfortunately hard is not defined it is quite possible that hard may refer to intact rock strengths in excess of about 50 to 75 mpa the rock mass classification approach is ideally suited for situations where the failure process is controlled by sliding and rotation of intact rock pieces this approach is less reliable for squeezing swelling spalling or slabbing ground or for rock bursting under very high stresses marinos et al 2005 the rmr adds together scores for ucs rqd joint spacing and condition and groundwater to get a value out of 100 table 8 4 6 the precision of the system is in the order of 10 units as reflected in the definition of five grades of rock the most common application is in the identification of stand up times a related system has been developed for coal measures molinda and mark 1994 there is also an application of rmr to slope stability romana 1993 geological strength index a relatively recent use of the rock mass classification systems particularly the rmr has been to use it to quantify the parameters required by a chile assumption for a rock mass hoek and brown 1997 since its introduction this approach has become very popular and tends to be used without adequate consideration of its limitations first the scale of the excavation must be such that the consideration is either of the intact rock itself or of a very large volume with sufficient discontinuities such that an equivalent continuum is present the implication of the intact rock scale is that one is dealing with a situation near the excavation boundary where the brittle parameters m 0 0 s 0 11 are more valid at the other end of the scale there is a need to ensure that throughgoing geological structures figure 8 4 1 are not present so as not to dominate behavior and make the continuum assumption invalid the method of determining the geological strength index gsi has been summarized in a number of cross plots e g figure 8 4 8 where consideration of the discontinuities is educed to one factor referred to as the surface condition th

e various plots suggest that a precision of 2 5 points may be the best possible this is assessed to be highly optimistic a precision of 5 is likely to be the best and a precision of 10 is a more typical result for rmr values in excess of 18 and 23 respectively the gsi is numerically equal to the rmr value if the latter is calculated with the 1976 version or equal to the 1989 version less 5 points an indication of the sensitivity of the key strength and deformation parameters to the gsi value is given in figure 8 4 9 the cohesion and tensile and unconfined compressive strengths drop rapidly with gsis between 100 and 70 depending on the gsi value the precision of the estimates of the various parameters is in the range of 10 to 50 a number of cautions apply marinos et al 2005 firstly there is the question of scale and the applicability of the mass criterion as discussed earlier secondly the method gives a non zero tensile strength that may not be justifiable in the case of discontinuous rock masses thirdly the surface condition is independent of the ucs of the rock which is counter to the jrc concepts for shear strength it is possible that this is implicit in some of the other cross plots for example the one for flysch that is included in roclab a software program from rocscience inc for determining rock mass strength parameters based on the generalized hoek brown failure criteria flysch is a sedimentary rock deposited in the marine environment it is possibly a soft rock it may be necessary to reduce the gsi by about 15 units if it is being applied to soft rock and finally there are concerns about the disturbance factor the depth of blasting damage may be in the order of a few meters for most blasts at this scale which is that of a roadway or a bench the hoek brown criterion probably should not be applied deformation parameters a number of analyses require values for the deformation modulus of rock masses as well as their strength interaction with water there are both chemical and physical deteriorations of rock mass strength with the presence of water the presence of water especially when combined with oxygen introduced by exposure to the atmosphere can lead to major changes in rock strength and usually strength reductions water pressures in the immediate vicinity of mine excavations are often but not always reduced to zero either because free flowing water is a nuisance underground or because drainage is critical to slope stability in the surface mines because of this it is conventional in mining rock mechanics to consider stresses in terms of total applied stresses there are some applications where the principle of effective stress needs to be considered effective stress equals the total stress minus the pore pressure in the context of shear along discontinuities the pore pressure can reduce the normal stress and hence it can reduce the shear resistance this is important behavior in

rock slope stability in addition low modulus rocks that are also aquifers may undergo significant deformations during depressurization such that the stress field is altered immediately ahead of the excavation compared to that of the far field condition this effect has been identified in porous sandstones and possibly in coal seedsman 2004 stresses rock and rock masses deform and collapse in response to selfweight and applied stresses self weight is the stress driver for planar and wedge failures and is ultimately the driver for most collapse modes there are some closed form solutions for the redistribution of applied stress around simple excavations since the 1980s and building on the availability of personal computers there have been huge advances in the scope and capability of computational methods to study the distribution of stresses around excavations in both two and three dimensions with the important proviso that the material can be considered to be linear elastic boundary element finite element distinct element and finite difference are merely different computational methods to solve the same complex equations and each has strengths and weaknesses in contrast to civil excavations mining ventures seek to operate at the limit of elastic behavior if not beyond so long as the workplace is safe and the business risks are acceptable because of the inability to adequately characterize the postpeak behavior of rock many mining problems are beyond the capacity of even the most complex numerical codes computational methods require knowledge of the stress field that was acting prior to mining there are now several tools to measure or estimate this stress field the stress field that is measured is itself a function of the properties of the rock mass in the immediate vicinity to which it is measured it is unlikely that the stress field can be adequately characterized prior to mining close observations of early excavations are still required and it is essential that the interpretation of the stress field does not become distorted by the models that are applied most observations are of deformations and this means that any calibration of the stress models includes assumptions not only on the deformation properties of the rock but also the stress redistribution about the excavation nature of stresses and stress equations the rock in which mining is conducted is stressed by forces associated with the gravitational weight of the overlying rock tectonic stresses and residual stresses related to earlier changes in depth of burial and rock temperature in mining rock mechanics it is conventional practice to assume that pore pressures are zero close to the excavation soil mechanics is the field of engineering that concentrates on the implications of non zero pore pressures and while this subdivision is convenient there are still many mining circumstances such as with very weak and extremely weak rock where the concepts of

effective stress need to be invoked any stress field is 3 d and it is conventional to resolve the field into the three principal planes and their associated principal stresses principal planes are those planes on which there are no shear stresses acting and by definition the principal stresses act normal to these planes the faces of excavations and the earth s surface are free of shear stresses and hence they are locally principal planes in situ stresses the world stress map heidbach et al 2008 can be used to obtain an initial assessment of the direction of the major principal stress in a horizontal direction a better estimate may be obtained if the structural geology of the deposit is known the local stress field is highly likely to be aligned with one of the axes normal to any dominant throughgoing discontinuity in fact with knowledge of the fault orientations and the associated fiction angles it is possible to constrain the estimate of the stress field in layered materials close to the surface e g within 200 m one of the horizontal stress axes must be parallel to the strike of the layers if it develops borehole breakout is a reliable indicator of stress directions that can now be readily determined using downhole tools many researchers have made compilations of stress magnitudes with depth until more site specific information is available it is reasonable to assume as a first approximation that the vertical stress can be calculated from the depth of cover and the major principal stress has a magnitude of between 0 3 and 2 5 times this vertical stress to a depth of 1 000 m and a magnitude of between 0 3 and 1 0 times this vertical stress below 1 000 m it is emphasized that the stress magnitudes are not the result of the so called poisson s ratio effect if stresses were considered to be applied in plane strain then higher stress magnitudes would be found in the stiffer units the stress field in coal is different to the stress field in the adjacent rock part of this may be due to the stiffness contrast but there is also a component that may be related to stress redistributions associated with volume changes as the coal is depressurized by mining stress magnitudes near faults will be less than at distant locations it is apparent that predicting the stress field before mining is difficult there are several techniques for measuring the stresses prior to mining but they are relatively expensive and because of the same geological controls that make the stress field variable it is possible that they will not be representative of the whole mine until in mine observations are possible the approach should be to combine any site data with inference of stress directions from an interpretation of the structural geology and to consider a range of stress ratios between 0 3 and 2 5 it is important not to ignore the possibility of low ratios such as for example in coal mining elastic redistributions the ready availabil

ity of numerical codes makes the calculation of stresses around excavations in two and three dimensions a relatively trivial exercise providing that the material can be considered linear elastic as the graphics are improved the interpretation and understanding of the stress field in three dimensions is improving importantly in linear elastic theory the redistribution of stresses is a function of the stress ratio and the aspect ratio and independent of the deformation properties and this applies to all excavation shapes large excavations are less stable as a result of the greater likelihood of encountering discontinuities for circular openings k values less than 0 33 induce tensile stresses in the crown and for a w h value of 2 a k value of 0 5 induces tensile stress in the crown recalling that a rock mass with discontinuities should be assumed to have zero tensile strength the importance of induced tensile stresses cannot be overestimated the onset of tensile stress should be considered to be the precursor to possible collapse that may develop rapidly and without warning the two equations suggest that there is an ability to modify the induced stresses by modifying the aspect ratio of an excavation the maximum boundary stress can be reduced by having an aspect ratio similar to the stress ratio it is unlikely that a mining operation will have such flexibility in coal mines continuous miners result in rectangular roadways and if the seam is flat lying two of the stress axes will be co planar with the bedding typical elastic stress patterns in isotropic materials are shown in figure 8 4 11 for a 6 1 5 m rectangular opening and for a circular opening for a 2 1 stress ratio the following three stress components are of interest 1 vertical stress syy elevated stress can cause failure in the rib 2 horizontal stress sxx low values could allow mobilization of vertical joints 3 deviatoric stress s1 s3 high values could cause compressive failure of the rock for the case where k 2 0 the vertical stress at the upper corner of the excavation is increased by about 14 and at the midpoint of the sides the vertical stress is reduced to about 60 of the initial value the horizontal stress is increased by about 60 at the corner and reduced to 80 at the centerline of the roof the deviatoric stress increases fourfold at the corners and doubles at the centerline of the roof another characteristic is the tendency for the deviatoric stresses to arch over rectangular excavations whereas they concentrate in the crown of a circular excavation figure 8 4 11a and 8 4 11b shows the location of three points that are located 250 mm in from the boundary at the spring line and the crown and also 250 mm in the roof directly above the rib line the relationships between the stress magnitudes at these three points with variations in the stress ratio and the aspect ratio of the roadway are shown in figure 8 4 12 the following key obs

ervations can be made the horizontal stresses in the roof become tensile at k values of about 0 6 the vertical stresses are higher and the horizontal stresses are lower for the excavation with the flatter aspect ratio the deviatoric stresses for the rectangular roadways are between 2 to 3 5 times the vertical stress the horizontal stresses at the crown of the circle are greater than at the roof centerline of the rectangular excavations for layered materials the isotropic assumption can result in the anticipation of lower stresses than actually develop for a condition where the value of independent shear modulus g is 1 100 of the isotropic value the induced deviatoric stress within 0 5 m of the excavation boundary may be higher by approximately 50 for a rectangular excavation and 100 for a circular excavation the magnitude of the induced stresses decrease with the square of the distance from the excavation so that at two diameters away the stress change is less than 5 these simple elastic models readily demonstrate that large excavations can induce major changes in the stress field around adjacent smaller excavations this is well recognized in metaliferous mining and is also a feature of longwall and pillar extraction in coal mines the extraction of a cut and fill stope can result in temporary increases in stresses as the excavation approaches the level of access drives and then a relaxation of stresses as the extraction level rises figure 8 4 13 in a longwall both the horizontal and vertical stresses increase at the face corner against unmined coal in the tailgate adjacent to previously extracted coal the horizontal stresses decrease and the vertical stresses increase an example of parametric analyses of the deviatoric stresses that can develop at the base of surface excavations is summarized in figure 8 4 14 in these analyses the stresses are generated by self weight only and the figure presents the normalized stress magnitude along a line drawn vertically down at the toe the normalization is based on the deviatoric stress that was present at the toe horizon prior to excavation depending on the face angle the deviatoric stress can be concentrated two to three times and the zone of influence is restricted to about 30 of the excavation height nonelastic behavior around excavations yielding yielding can result from either the peak strength of the rock being exceeded or shear along discontinuities being induced either way movement is induced in most mining situations the applied stresses can be considered to be stiff such that they are redistributed elsewhere if the rock about the excavation moves beyond the elastic limit figure 8 4 15 the limit of elastic deformation is in the order of 10 mm the implication is that after initial roadway formation the stresses in the immediate roof should be assumed to be very low this behavior and the limitation of many computational methods to adequately mana

ge it is particularly important when roadways experience changes in stress conditions during their lifetime stress relaxation can also occur in the roof if there is yielding or failure in the sides or floor of an excavation figure 8 4 15 in this case the relaxation may be sufficient to induce the onset of tensile horizontal stresses and the mobilization of steep or vertically oriented joints the amount of deformation that is required for relaxation of roof will be partly a function of the relaxation that has already developed in the roof the hazard should be anticipated if the roadway defines a pillar that is designed to yield or if there is a lowstrength floor horizon that may compress or fail body stresses recalling that rock masses are assemblages of blocks if the blocks begin to rotate they can interact and generate areas of high compressive stress while in immediately adjacent areas there can be an opening of joints in response to localized tensile stress an example of this is the voussoir beam mechanism sofianos and kapenis 1998 that develops in a roof with no applied compressive stresses and yet failure can be via compression figure 8 4 16 cantilevering behavior can also be anticipated in some layered materials both these two behaviors may generate compressive failure in the roof rib corner at the same location as stress concentrations develop in linearelastic codes an implication of this is the need for an underground observer to be careful about inferring the nature of the applied stress field from simple observations an analysis of likely stress magnitudes and rock strength is needed before the stress field can be inferred practical implementation of rock mechanics in many respects this chapter highlights the many uncertainties that are inherent in the application of rock mechanics to mining the general nature of these uncertainties is no different to those that face any engineering venture in soil or rock absolute guarantees of acceptable performance are not possible the rock mechanics engineering profession has developed a number of strategies to manage the hazards and risks that are generated by the uncertainties of rock observations and monitoring are recognized as an essential part of developing and applying empirical reasoning to any mine design recognizing the uncertainties of soil and rock the observational method peck 1969 was introduced as a formalized process to manage contractual risks and to avoid latent condition claims it requires full characterization of rock mass and prior assessment of likely and extreme conditions that may be encountered such that pre agreed responses are available it is stressed that monitoring is not an alternative to fundamental design because it requires detailed knowledge of conditions the observational method should only be considered for specific well constrained projects both in time and geographically it is obvious that the method cannot be applied to dec

isions of major capital expenditures prior to the operational phase bulk mining systems require very large daily production and any unscheduled delays can have major cost impacts it is possible that bulk mining systems should not consider the observational method as they can afford demand greater geotechnical certainty that can be delivered by a worst case design monitoring to calibrate a complex numerical model may mean that in the initial stage of operations there is no acceptable design in place careful consideration of the safety and business implications of this is needed monitoring may also be required for triggered action response plans tarps but these plans are principally about personnel safety with a secondary application of stepping between agreed support strategies as in the observational method tarps are an essential part of modern mining as they provide the rationale to evacuate a place but they are not sufficient to ensure an adequate mining outcome if there is inadequate time to stabilize the ground an overreliance on the observational method and or inmine monitoring to validate designs can lead to poor business performance what is needed is a rigorous design process for rock bieniawski 1993 the following are key aspects of the design methodology figure 8 4 17 safety of the work force is the key requirement there may be practical constraints to the ideal solution as much geological data as possible should be collected and interrogated numerous pathways are available figure 8 4 2 and the choice of pathway must be justified uncertainties in the geological input requires sensitivity studies on the outcomes a number of options with different levels of business risk may be identified the final recommendations may carry residual business risk which needs to be highlighted monitoring to confirm design assumptions may be additional and to separate from that required for tarps the general methodology can be applied at all stages of a mining project with different pathways more applicable at different stages figure 8 4 3 even within a single pathway there may be differences in how the analyses are conducted and what behavior models are used the underlying principle is to maintain flexibility and be prepared to apply judgment if the analyses do not appear sensible the state of the art is still not perfect overall mine design because modern mining requires large capital investments it is not valid to advocate the observational method for overall mine design there must be a high level of confidence that the capital can be recovered within the indicated time line a more conservative mine plan may be required initially until the uncertainties are understood once the capital is recovered the focus can then move to maximizing longterm revenue there are adequate tools to design the general size and shape of excavations in a mine computational methods in two and three dimensions using equi

valent continuum assumptions and the generalized hoek brown criterion are appropriate for underground metal mines the simpler block geometry of many underground coal mines and bench scale designs in surface mines allow for the application of simple behavior models and limit equilibrium analyses specific analyses of major throughgoing geological structures need to be considered as early as possible the stress field assumptions that may be necessary prior to mining need to recognize the uncertainties in their measurement and how well they represent the entire deposit a check of the stress field against static equilibrium on pervasive weakness planes should be conducted if possible decisions on the orientation of the excavations should be deferred as long as possible so that in mine knowledge can be obtained failure and collapse modes around excavations the serviceability of individual benches in surface mines and access roadways in underground mines is just as important as the performance of the larger extraction volumes this serviceability can be adversely impacted by failure of a component of the overall mine design such as a pillar or by the localized collapse of the side or roof of the roadway or an individual bench a logical framework for the design of specific excavations step 5 in figure 8 4 17 is presented in figure 8 4 18 and is equally applicable to surface and underground excavations with slight changes in emphasis the rock mass classification pathway using the rmr or q systems is an alternative to the use of this flow chart in underground excavations the first step is to determine the boundary stresses this is now a relatively trivial task with a range of software programs available for chile materials if the lithologies are layered there may be a need to consider the sensitivity of the stresses to the assumption of isotropic behavior determining the stresses at the excavation boundary induced by subsequent stress changes is still a challenge for surface mines the assumption of body forces only is valid for routine analyses any consideration of the applied stresses in a surface mining application is best left to specialist geotechnical analyses in which the progressive excavation is modeled roof when selecting the discontinuity branch in figure 8 4 18 there are readily available limit equilibrium methods to assess the stability of joint bounded blocks and to specify support strategies accepting the concept that bedding partings define beams there are also a number of simple elastic beam and voussoir beam models that can be applied the proximity of the roof bolting to the mining face will determine if the determination of the forces of slip and separation need to be estimated with 2 or 3 d stress models if bolting is conducted close to the advancing mining face within a distance equal to the span a 3 d model will give a better estimate of the forces developed after the support is installed there are many

situations where the elastic boundary stress may be tensile it is good practice to assume that the rock mass will have a tensile strength of zero so the possibility of tensile elastic boundary stresses should be carefully reviewed as failure is indicated depending on the shape of the excavation tensile conditions will be encountered if the k ratio is less than about 0 5 i e when the vertical stresses are significantly higher than the horizontal stresses low k ratios should be anticipated in roadways close to the sides of valleys or surface excavations when large excavations are in close proximity and in coal roofs because there is no new rock breakage just relaxation across joints tensile failures in rock and coal masses can occur without audible or visual warning the onset of compressive failure can be readily assessed if the brittle failure criteria are considered as the procedure is to compare the deviatoric stress with one strength parameter ucs a particularly useful step is to normalize the ucs to the estimated in situ vertical stress based on the depth of cover and the average density depending on the k ratio and the shape of the excavation values less than about 4 can be interpreted to indicate a greater likelihood of compressive failure at the excavation boundary sides pit walls and underground on the bench and access roadway scale the discontinuity branch in figure 8 4 18 leads to the analysis of planar slides wedges and topples rotational slips should also be considered numerous solutions and associated software are available the brittle failure criterion is also applicable to the sides of underground excavations in fact the identification of this failure criterion was based on extensive work on the spalling of the sides of roadways in massive rock floors adverse behavior of the floor should also be considered assuming the same deformation properties the elastic deformation of the floor should be of a similar magnitude as for the roof in the order of 10 mm because of the highly likely presence of water in mine floor there may be some swelling of the contained minerals and some additional floor movements buckling of thin floor beams may also need to be considered floor heave is a relatively common occurrence even in hard rock mines under conditions of high stress in the underground coal sector the possibility of a bearing capacity failure along thin layers of low strength clayey materials needs to be considered this is better considered in the context of soil mechanics conclusions rock mechanics knowledge continues to evolve rapidly the basic science is well established and advances in computer analyses provide much greater insight into the complex stresses that develop around and between mining openings rock masses are complex and their reaction to changing stress conditions is controlled by the inherent strength of the rock material movement along large or small scale discontinuities

and the presence of water the application of rock mechanics continues to drive improvements in safety and mine productivity and even greater success will be achieved as the ability to characterize and quantify rock mass strength improves ground control practices in underground mines are constantly evolving always with the goal of improving mine safety and economics these changes are often influenced by the mining methods used the level of mechanization and types of equipment used and the labor cost and skill set ground control practices are further defined by rules and regulations linked to risk tolerance and to the mining cultures of different jurisdictions or countries in this chapter a distinction is made between reinforcement and surface support reinforcement is a technique in which elements such as rock bolts are applied internally to the rock mass chapter 8 8 provides a thorough review of reinforcement techniques surface support is a technique in which elements such as shotcrete or steel mesh the term screen is also commonly used in north america are applied to excavation surfaces externally to the rock mass this chapter examines surface support as it is used for underground metal mining a comprehensive collection of surface support applications in mines is presented in potvin et al 2004 the traditional approach to ground control in underground metal mines was to install reinforcement to stabilize potential wedges and blocks that might daylight at the surface of excavations this required good ground awareness skills on the part of the mine workers who used their experience to read the ground and install reinforcement at appropriate locations to stabilize the rock mass around excavations small blocks or scats that formed at the rock surface around the excavations due to stress changes and blast vibrations were controlled by frequent scaling when ground conditions were less competent the rock surface was controlled with mesh the main role of surface support was to support the dead weight of smaller rock blocks between the rock bolts since about 2000 a concerted effort toward minimizing the risk of rockfall injuries has accelerated the development of improved ground control practices when it was recognized that small rock falls cause many injuries the systematic use of surface support became common practice in most mechanized mines nedin and potvin 2005 several mines changed their ground support strategy designed originally to suit local conditions to the systematic application of pattern bolting and surface support current ground support standards are well documented and reflect the current industry s low risktolerance approach to rock falls in this new ground support strategy reinforcement elements hold or pin larger blocks and surface support elements catch smaller blocks however this approach is less applicable in ground conditions where the rock mass is already in a postfailure or yielding

state due to excessive static or dynamic loading in these cases the excavation often experiences large displacements and the surrounding rock mass crumbles to maintain the integrity of the excavation surface wall or roof surface support should be capable of containing the bulk of the broken rock mass and later movement resulting from stress changes such as stoping the surface support must deform with the rock mass while the load transfers from the surface support through the plate or surface fixture arrangements to the reinforcement the function of reinforcement is to retain the surface and limit displacement or convergence thus reinforcement and surface support are part of an integrated system that transfers and shares load until the excavation surface is stabilized or until the support system fails at its weakest link surface support elements although a wide variety of rock bolts are available the choice of surface support elements is limited to steel mesh straps and shotcrete steel mesh is the most popular in metal mines but shotcrete has been gaining in popularity since its introduction in the late 1980s straps also considered to be surface support even though they do not completely cover the rock surface are used to contain the rock mass between reinforcement elements and distribute load between the reinforcement elements used to install the straps steel mesh two main types of steel mesh are used in the mining industry welded wire mesh weld mesh and chain link mesh also called diamond mesh expanded metal mesh is also available but is not widely used mines that use a mechanized approach to mesh installation use predominantly precut sheets of welded wire mesh welded wire mesh sheets of welded wire mesh are precut to customized dimensions in australia sheets are generally large 2 4 3 0 m 8 10 ft to promote productivity accounting for overlap a row of six sheets is commonly used to cover 40 m2 130 ft2 in north america installation involves more manual handling and sheets tend to be smaller 1 2 1 5 3 0 m 4 5 10 ft and thus lighter in weight typical dimensions for weldedwire mesh are shown in figure 8 6 1 notwithstanding quality control issues and assuming the use of standard aperture welded wire mesh wire thickness t in figure 8 6 1 is the main influence on wire performance in north america wire thickness is expressed in terms of gauge 4 6 and 9 gauge 5 8 4 9 and 3 7 3 8 mm diameters respectively are most commonly used in australia a diameter of 5 6 mm is most commonly used but sometimes diameters as small as 4 95 mm are also used selection of mesh aperture s in figure 8 6 1 is in theory based on the smallest block that the mesh must contain in north america mesh aperture is often standardized at 100 100 mm 4 4 in a smaller aperture such as 50 50 mm or 75 75 mm 2 2 in or 3 3 in is used where extra load bearing capacity is required a larger ap

erture such as 150 150 mm 6 6 in can be used when shotcrete is to be sprayed over the mesh the larger aperture allows for good penetration of the shotcrete through the mesh when the mesh is expected to provide support for a long time or in corrosive environments galvanized coating is used to protect against corrosion welded wire mesh is relatively inexpensive and can be installed quickly because it is easily attached to existing reinforcement by means of plates and rock bolts it can support small blocks and its energy absorption properties enable it to provide some control during rock bursts it can also be used with shotcrete no 9 gauge mesh is susceptible to damage from fly rock no 6 gauge mesh can be a good substitute but is less flexible and more difficult to install around corners mesh is fabricated using longitudinal wire to which cross wires are spot welded when it is to be installed near an advancing face optimally the cross wire should be laid against the rock surface and the longitudinal wire laid on the outside facing the opening and along the drift axis this procedure mitigates blast damage by flying rocks which are less likely to catch into cross wires in north america it is popular to install ground support systems using dedicated bolting machines the maclean bolter features a scissor platform with a rock bolting boom mounted at the tail end and operator controls mounted at the front end figure 8 6 2 hooks to hold sheets of mesh are located on the side of the platform installation of bolts and mesh in a single pass requires some manual handling of the mesh during bolting mesh can also be installed in a second pass after bolting is completed by means of pressure plates fitted over installed threaded bars figure 8 6 3 welded wire mesh is commonly installed manually by means of an air leg which is often the only option in countries where mining mechanization is low manual installation is performed with a handheld drill in smaller excavations or from scissor platform trucks or integrated tool carrier baskets in larger drifts in practice it is difficult especially with largediameter wire or stiff welded wire mesh to ensure that the mesh sheets are installed tightly against the rock mass surface and manual installation increases the risk of minor injuries a number of other quality control issues are common to all welded wire mesh installations in particular poor overlap between sheets is highly undesirable because it may later expose mine personnel to rockfall hazards standard practice is to overlap three squares of mesh but in difficult ground conditions drift profiles are often erratic and proper overlap can be difficult to obtain villaescusa 2004 identified three modes of mesh failure 1 shear failure at the weld point presumably due to the technology used for welding and to quality control issues such as dirty electrodes or dirty wire 2 heat affected zone failure due to excess

ive weld head pressure and temperature 3 tensile failure of the wire due to excessive tensile load the load and displacement capacity of mesh depend on the type of metal and size of the wire and dictate the load bearing capacity of the wire and weld it follows that weld capacity should be comparable to wire strength tannant 2004 and thompson 2004 developed testing procedures to better define the capacity of mesh and published pull test results for a variety of mesh and bolt arrangements on tests of standard 5 6 mm diameter australian weldedwire mesh thompson found load capacity to be 20 40 kn with displacements of 100 350 mm 4 14 in the variability of results arises from variations in simulated bolt patterns and test procedures in general a smaller bolt pattern 1 1 m 3 3 3 3 ft supports higher loads but produces a stiffer response than does a larger bolt pattern 1 5 1 5 m 5 5 ft tannant tested load as a function of displacement for the three common mesh gauges used in canada in a simulated bolt pattern of 1 2 1 2 m 4 4 ft figure 8 6 4 chain link mesh chain link mesh is used less than welded wire mesh its main advantage is its energy absorption capacity which is associated with its greater displacement capacity consequently one application of chain link mesh is part of a rock burst support system its main disadvantages are that it tends to unravel when damaged and that it is difficult to apply shotcrete through it typical chain link mesh dimensions are shown in figure 8 6 5 chain link mesh differs from welded wire mesh in its ease of installation and its deformation capacity it has very low rigidity which makes handling and installation awkward and difficult it comes packaged in rolls and needs to be gradually and manually unrolled during installation to its final position against the rock surface installation is labor intensive and difficult to mechanize as a result it is not widely used in highly mechanized underground mines but remains popular in the mines of for example south africa and south america mechanization of chain link mesh installation is currently being addressed by development of the rock mesha mechanized mesh handler which can be mounted on one of the booms of a twin boom jumbo drill figure 8 6 6 coates et al 2009 one boom unrolls the mesh against the rock surface while the other boom installs bolts a prototype of this equipment is being tested in australia interestingly the low rigidity of chain link mesh can be an advantage in terms of its deformation capacity tannant 2004 pull tested three types of mesh chain link weldedwire and expanded metal figure 8 6 7 and table 8 6 1 expanded metal mesh is seldom used in underground mines and is not discussed here chain link mesh demonstrated a displacement capacity of 400 mm 16 in at peak load with good post peak behavior retaining 50 of its load capacity at 800 mm 32 in displacement the load capac

ity of chainlink mesh like that of welded wire mesh is largely a function of wire size and the type of steel used in canada 9 gauge 3 7 mm wires are commonly used in south africa wires of 3 2 4 mm are standard high tensile light steel chain link mesh is also available for use when ground conditions require extra load bearing capacity whether the situation is static or dynamic coates et al 2009 for example high tensile tecco mesh which uses wire 4 mm in diameter and weighs only 2 6 kg m2 0 53 lb ft2 was tested at 110 kn using the test procedure described in player et al 2008 load transfer between mesh and reinforcement the load bearing capacity of mesh is well suited for containing small blocks that can detach between reinforcement elements standard reinforcement patterns used in mines generally call for elements to be spaced 1 1 5 m 3 3 5 ft apart in both directions so that the maximum block weight that can fall from between the bolts is 2 t or metric tons 20 kn well within the capacity of most mesh the detached blocks exert a pulling force which is transmitted from the mesh to the bolts via the bolt plate the role of the bolt plate is to transfer force to the mesh maintain tension in the bolts keep the mesh and straps in place and maintain the confinement of rock blocks beneath the collar when bolt plates are small sharp and square the total force may be transmitted entirely through only a few wires and the sharp edge of a plate can cut the wire this situation can be remedied by using larger plates to distribute load to more wires or by using shaped plates to minimize cutting edges circular plates are available as are other specially designed plates such as butterfly plates which are large thin and deformable the more popular bolt plates are shown in figure 8 6 8 and their dimensions are listed in table 8 6 2 it is generally accepted that domed plates provide the best contact for irregular rock surfaces if the rock surface is very irregular more flexible plates 6 mm 0 25 in should be used the weakest link of the total support system can be the mesh the plate the connection between the plate and the mesh e g due to plates cutting wires or the rock bolt rock bolts and plates are designed to fail at comparable loads so as to avoid a weak link however this is true only under ideal tensile conditions when the load has a strong shearing component or a point load is applied to one component because of incorrect installation early failure of the bolt heads or the plates can occur the heads ring of friction bolts are particularly prone to early failure when bolts are installed at an angle to the rock surface or when they are hammered in too hard in squeezing ground or in rock burst conditions where the rock mass experiences large deformation both the load capacity and the deformation capacity of the support system are tested comparison of the load and deformation capacities

of mesh and bolts shows a clear mismatch in their properties grouted bolts are stiff and allow displacements of only a few millimeters before failure but they have a load capacity in the range 150 200 kn in contrast mesh allows large displacements of several hundreds of millimeters but has a load capacity of only 20 40 kn as the ground deforms mesh deforms with it but fails when either the local load exceeds 20 40 kn or the displacement exceeds 100 350 mm 4 14 in in these cases mesh is incapable of transmitting loads that exceed its own load or displacement capacity to the bolts a tighter bolting pattern in such cases divides the excavation surface into smaller areas between the bolts resulting in lower local loads for the mesh to contain where ground displacement is considerable bleeding the mesh and rehabilitating the ground support is a recurring issue straps straps are generally used as surface support in special applications rather than as a part of a systematic support system in mines they can be installed either on their own against the rock mass surface by means of rock bolts or over mesh to provide extra strength and stiffness to the surface support when used alone they can support key blocks between rock bolts however if key blocks behind the straps loosen and fall the strap no longer has good contact with the rock surface and may become ineffective in this area best results are obtained when the straps are installed tightly across the most prominent structures common practice is to wrap straps around pillar noses to prevent progressive deterioration of the pillars particularly those submitted to vertical load the three types of strap used in mining operations are steel straps mesh straps and osro straps shotcrete shotcrete is concrete that is applied to the rock surface by pneumatic shooting at high velocity using specialized equipment it is composed of 15 20 cement 30 40 coarse aggregates 40 50 fine aggregates and 2 5 additives water is added to the mix to hydrate the cement and initiate the chemical reaction whereby the concrete gains strength the water must be clean and free of chemical components that are potentially detrimental to the reaction the water cement ratio expressed in units of liters per kilogram influences cement hydration the final product should be in the range 0 3 0 5 l kg often steel or synthetic fibers are added to the mix to improve the post peak behavior of the concrete the resulting product is commonly called fiber reinforced shotcrete frs or fibercrete the use of shotcrete as a surface support as well as for construction applications in mines has increased rapidly since the 1980s the flexibility of the equipment the choice of mixing methods and the availability of shotcrete has contributed to its growing popularity as the preferred surface support method in underground mines shotcrete is often perceived as the surface support method of choice when the

rock mass is heavily fractured its initial adhesion properties lend it a gluelike effect on heavily broken rock mass which helps prevent the rock mass from crumbling contrary to the case for mesh which contains blocks of rock only after they detach from the surface and therefore begins its action after the rock mass has failed shotcrete provides immediate reactive support action at very low displacement as soon as the rock mass starts to move and long before it reaches a yielding state as a result shotcrete is much more efficient at preserving the confinement within the rock mass surrounding an excavation in recent years systematic application of shotcrete after each blast also known as in cycle shotcreting has gained popularity as an alternative to the so called campaign shotcreting of an entire section of tunnel in cycle shotcreting has a number of advantages and disadvantages its advantages are as follows 1 its rapid application minimizes time dependent rock mass deterioration 2 it allows for reinforcement to be installed after the shotcrete by means of bolts that pin the shotcrete to the wall and 3 it favors good load transfer between the surface support and the reinforcement a disadvantage is that the rock mass is covered immediately after being exposed leaving little opportunity for geological and geotechnical data to be collected new data collection techniques such as digital stereophotogrammetry are being used to overcome this problem two main techniques exist for applying shotcrete in mines dry mix and wet mix in the dry mix technique water is added to the mix only at the nozzle advantages and disadvantages of the technique are as follows based on aci committee 506 2005 and the authors experience advantages of the dry mix technique maximum flexibility in delivery in terms of time because curing begins only when the shotcrete is applied flexibility in terms of transportation because shotcrete ingredients can be carried even long distances to the workplace in premixed bulk bags and the necessary equipment including mixers pumps hoses and more is compact and mobile this flexibility makes this method very popular in deep mines that are accessible only by shaft instantaneous control over mixing water and consistency of the mix at the nozzle to meet variable field conditions high suitability for placing mixtures containing lightweight aggregates or refractory materials disadvantages of the dry mix technique large percentage of rebound dusty environment created low application productivity low volume per hose size in the wet mix technique shotcrete mix is prepared at a custom built concrete plant often located on the surface but sometimes underground or from a local supplier shotcrete is transported by agitator truck from the plant to the point of application alternatively a slick line can be installed to alleviate some of the transportation difficulties of bringing

underground shotcrete and other concrete used for construction advantages and disadvantages of the technique are as follows based on aci committee 506 2005 and the authors experience advantages of the wet mix technique generally minimal quality control issues because mixing is done in a controlled plant environment relatively low sensitivity to the skills of the nozzle operator high delivery rates low percentage of rebound minimally dusty environment ease of mechanization control of mixing water at the delivery equipment where it can be accurately measured good assurance that mixing water mixes thoroughly with other ingredients minimal dust and cementitious materials lost during the shooting operation normally low rebound resulting in less waste high volume per hose size disadvantages of the wet mix technique transport and delivery logistics are difficult in deep mines having shaft access only equipment is more expensive not suitable for small volume applications a prerequisite for the successful application of shotcrete is the use of properly operated and maintained equipment by qualified personnel in general dry mix equipment consists of either single or double chamber guns batch or rotary continuous feed guns with batch guns a charge of material is placed in the chamber and the chamber is closed and pressurized feeding the material into a delivery pipe or hose with double chamber guns the upper chamber serves as an airlock during the charging cycle to allow for continuous operation rotary guns use the rotating airlock principle to supply a continuous feeding action wet mix applications make use of pneumatic feed guns slugs of material are introduced into the delivery hose and compressed air is added at the discharge sump at the nozzle to increase the velocity of the mixture positive displacement guns can also be used a solid column of shotcrete is forced by mechanical air or hydraulic pressure through a hose in a continuous stream to the nozzle air is injected at the nozzle to break up the stream and increase exit velocity basic components of shotcrete the main components of shotcrete cement aggregates water admixtures and potential cementitious additions are described in vlietstra 2009 and aci committee 506 2005 cement the binding agent in shotcrete is cement quite often an ordinary portland cement portland cement contains lime silica alumina and iron oxide lime is the main component the other three are included to balance the chemical composition of the cement to control the initial reaction of the cement with water the hydration process gypsum is also added the chemical reaction of hydration is complex and beyond the scope of this chapter it is important to note however that the process is time dependent being slower during the first few hours it is also temperature dependent being faster at higher temperatures the other factor affecting the hydration

process is grind size which is often measured in terms of surface area and quantity of 45 m residue finer grind facilitates more rapid hydration and earlier strength gains aggregates aggregates form the matrix of the shotcrete and are bound together by a cementitious paste to become concrete both coarse and fine aggregates are used the combined aggregate should generally comply with one of the gradations listed in table 8 6 3 maximum aggregate size is dictated by pumping limitations and the need to minimize rebound coarse particles can block the nozzle and subsequent cleaning can be timeconsuming larger particles tend to rebound when sprayed on a hard surface they can also penetrate already placed shotcrete and produce craters that are difficult to fill melbye et al 2001 a guideline to avoid segregation and clogging is to ensure that 30 passes through a 0 3 mm sieve if the proportion of fine material exceeds this more water is required for hydration aggregates should be clean that is free of dirt chemical impurities and organic material the mineralogical composition of the aggregate should be inert under normal mining conditions and free of sulfides and other reactive minerals equidimensional aggregate particles generally give better results than do those with elongated or platy shapes and a rough surface assists cement paste bonding aggregates must be at least as strong as the specified concrete final product water mixing and curing water should be clean and free of material that can be detrimental to concrete if potable water is not available the water must be tested to ensure that the strength of mortar cubes made with it is 90 of the strength of mortar tubes made with distilled water admixtures admixtures are chemicals that are put into concrete to affect the mixing placing and curing processes typical admixtures are accelerators air entraining agents and pozzolans for wet mix shotcrete a consistency corresponding to a target slump of 150 mm 6 in facilitates mixing pumping and shooting in a slump test a concrete sample is placed in a metal slump cone of height 300 mm 12 in bottom diameter of 200 mm 8 in and upper diameter of 100 mm 4 in the concrete is tapped down following astm c143 c143m 2009 procedures and the cone is lifted vertically to remove it without disturbing the sample the sample is allowed to slump until it stabilizes the slump height is the distance that the sample settles below its original 300 mm 12 in height because consistency and slump are controlled mainly by the water cement ratio which in turn influences the final strength of the shotcrete water reducer admixtures plasticizers and superplasticizers are often used to achieve the desired consistency for shotcrete application hydration controllers are also often required to delay hydration and accommodate the time required for the wet mix to be transported from the plant to the working face without ad

mixtures shotcrete shelf life would be only 1 to 2 hours depending on ambient temperature hydration controllers can extend shelf life up to 3 days without affecting the short term or long term strength of the shotcrete an accelerator can be added at the nozzle to cancel the retarding effect of the hydration controller for both dry mix and wet mix techniques accelerators are required to obtain early adhesion to the rock surface and induce rapid gains in shotcrete strength accelerators are also used to apply thick layers of shotcrete in a single pass unfortunately however accelerators tend to reduce the final strength of shotcrete accelerator dose is generally expressed as a percentage of the total weight of cementitious content doses are typically 2 5 for dry mix and 3 10 for wet mix a number of accelerators are available and performance testing is strongly recommended prior to product selection because local cement can respond differently to accelerators noncaustic alkali free accelerators are gaining in popularity because they are relatively environmentally friendly and induce rapid gains in shotcrete strength with minimal long term loss of strength cementitious additions it can be advantageous to replace a portion of the cement with pozzolanic material such as microsilicia silica fume or fly ash if these materials can be obtained at reasonable cost they can accrue economic benefits and improve some of the shortterm and long term properties of shotcrete silica fume is an ultrafine material particle diameter of 0 1 0 2 m whose spherical particles have high pozzolanic properties in the short term it has a lubricating effect on the mix increasing flowability and reducing equipment wear it also minimizes rebound and aids adhesion to rock in the long term it improves shotcrete strength and durability fly ash is a coarser material particle diameter of 50 m compared to silica fume it is generally less expensive but also less effective in improving shotcrete properties reinforced shotcrete concrete including shotcrete is a brittle material with high compressive strength but low tensile and flexural strength the failure mechanism for shotcrete is discussed in the shotcrete stabilization mechanisms section the way it is applied in underground mining causes it often to be subject to bending and flexural forces therefore some type of reinforcement is required in most mining applications two common types of reinforced shotcrete are used in mines mesh reinforced shotcrete and frs both are widely used and have a number of variations mesh reinforced shotcrete in the early days of shotcrete application in canadian mines in the 1980s it was common practice to apply shotcrete over mesh as a reactive action particularly when ground conditions were deteriorating the results were often very good and acceptance of the technique was rapid the technique then became proactively used as a high capacity surface suppo

rt system applied in areas where difficult conditions were anticipated mesh reinforced shotcrete remains widely used in canada and south africa in extreme conditions mine operators have tried to embed welded wire mesh inside shotcrete by placing a layer of shotcrete first installing mesh over this first layer and then placing a second layer of shotcrete the result is a thick shotcrete liner that can maintain a good proportion of its load deformation capacity after the shotcrete cracks however when deformation reaches 10 20 mm the wires in the mesh which are pointloaded by the shotcrete crack tend to fail in tension an alternative approach that takes advantage of the ductility of the mesh is to install it as the final layer over the shotcrete preferably frs then when the shotcrete cracks it is contained by the mesh which is then free to deform to its full capacity of 100 350 mm 4 14 in installing mesh over frs tends to be the preferred option in australian mines facing difficult ground conditions fiber reinforced shotcrete fiber reinforced shotcrete is also called frs or fibercrete the addition of steel or synthetic fibers improves the postfailure behavior of shotcrete when it is submitted to bending or flexural forces fibers have no effect until the ultimate strength of the shotcrete is reached and cracks start to propagate through the shotcrete layer the fibers then bridge and prevent further opening of the cracks so that a significant fraction of the loaddisplacement capacity of the yielding shotcrete is maintained figure 8 6 15 to fulfill their role shotcrete fibers must have both high tensile strength and a strong bond with the shotcrete a strong bond is achieved with hooks or cavities two factors particularly affect the bridging performance of fiber the first factor is the length diameter aspect ratio of the fiber in calculating this ratio the diameter can be taken as the equivalent diameter of a circle whose area is equal to the cross sectional area of the fiber a high aspect ratio contributes to better bridging the second factor is the fiber dosage greater numbers of fibers that can bridge cracks contribute to better bridging several types of steel and synthetic fiber are available figures 8 6 16 and 8 6 17 synthetic fibers are usually made of organic polymers that are small enough to permit even distribution in the sprayed mixture during mixing and spraying fibers with equivalent diameters of 0 3 mm are called macrofibers fibers with equivalent diameters of 0 3 mm are called microfibers typical macrofiber aspect ratios are in the range 40 65 for common fiber lengths of 19 to 50 mm 0 75 to 2 in steel fiber lengths are generally 38 mm 1 5 in synthetic microfiber lengths vary from 6 to 50 mm 0 25 to 2 in aci committee 506 2008 originally steel fiber was used in frs applications for both civil tunneling and mining recently synthetic fiber has been gaining in popularity in civi

l tunneling and in australian underground mines among the relative advantages of synthetic over steel fiber is its greater resistance to corrosion lower abrasiveness and significantly lighter weight lighter weight allows for a significantly higher dosage of fibers similar performances are obtained with one fourth the weight of synthetic fiber which is also twice the volume as compared with steel fiber the elastic modulus of synthetic fiber 3 400 4 800 mpa 500 000 700 000 psi can be 50 times less than that of steel fiber with more and stiffer fibers per unit volume synthetic fiber locks cracks earlier and transfers load to the concrete better promoting preferential formation of multiple local cracks rather than a single crack opening progressively the main disadvantage of synthetic fiber is that it floats and can plug the pumps of underground sumps or cause problems in flotation cells bernard 2004 investigated the creep behavior of both steel and synthetic fibers and concluded that they have similar resistance to load deformation creep is an issue when shotcrete cracks with steel fiber reinforced shotcrete if a crack develops there is potential for corrosion of the fibers and loss of ductility load energy continues to be absorbed but there is a risk of sudden rupture in contrast with synthetic fiberreinforced shotcrete development of cracks is not associated with corrosion if the load continues synthetic fibers stretch rather than suddenly fail and continue to absorb load energy surface preparation profile and shotcrete thickness good rock surface preparation before shotcrete application involves removing or scaling down loose rocks and cleaning the rock surface of dirt grease residues from blasting and diesel exhaust and more kuchta et al 2004 recommend the use of high pressure water jets for scaling down loose rocks in mine openings in order to maintain mine safety good preparation improves the adhesion strength of the shotcrete usually defined as the strength of the interface between two materials in this context between rock and shotcrete malmgren et al 2005 reported the results of an extensive failure mapping investigation in 7 km 4 4 mi of shotcreted drifts at the kiirunavaara mine in sweden shotcrete was found most likely to fail where the shotcrete layer is thin 20 mm and adhesion to the rock is poor to nonexistent the main factors that contribute to shotcrete failure are rock deformation and dynamic loads due to blasting windsor and thompson 1999 argued that traditional shotcrete layers of uniform thickness are suitable only for regular and smooth or irregular and evenly coated excavations they suggested that depending on the excavation profile different cover techniques can be used wood 1999 built upon this work developing a methodology for estimating the volume of shotcrete required for any particular project such estimates are of particular interest for operations in whi

ch shotcreting is contracted out and it is necessary to determine the volume of shotcrete based on the design profiles of excavations the estimated volume is based on roughness and rebound the roughness factor is determined by multiplying the rating that represents the irregularity of the surface of the ground and the rating that represents the quality of the drill and blast process table 8 6 5 in excellent ground conditions the final perimeter of the excavation is according to the design specifications good ground conditions are characterized by the presence of joints and slightly rough surfaces fair ground conditions are characterized by blocky ground and smooth surfaces finally in poor ground conditions with very blocky ground polished surfaces possible fault zones and wet ground the rock mass has a significant influence on the perimeter of the excavation and is assigned a higher ground condition factor the quality of the drill and blast process is characterized by the degree of overbreak a good drill and blast process defined by half barrels across the back and some additional overbreak beyond design would be assigned a drill and blast rating of 1 10 if this was observed in good ground conditions with a ground condition rating of 1 20 the resulting roughness factor would be 1 32 the roughness factor and the rebound factor can be used to calculate a specific volume factor for any particular mining project table 8 6 6 rebound varies during the course of shotcrete shooting it is high at the beginning of the shoot and decreases with increased shotcrete thickness a site specific chart can be developed to quantify cumulative rebound with respect to average thickness of shotcrete placed based on shotcrete trials this can be used to establish a rebound factor wood 1999 presents an example in which the design calls for a minimum shotcrete thickness of 25 mm 1 in so an average of 50 mm 2 in of shotcrete is probably needed to provide the desired cover associated with a minimum of 15 rebound similarly a minimum design thickness of 50 mm 2 in requires an average thickness of 75 mm 3 in and the rebound is lower resulting in a lower rebound factor the rebound factor can then be multiplied by the roughness factor obtained from table 8 6 5 to generate a site specific volume factor gaps in the table indicate situations that are unlikely to occur in practice for example it is unlikely that a rock mass with poor roughness would require only 50 mm 2 in of shotcrete similarly there is little justification for 100 mm 4 in of shotcrete in a high quality well excavated rock mass a practical consideration in rapid development headings is the time required after shotcrete application before it is safe to reenter under newly secured ground clements 2004 reviewed the following tests to determine in situ early strength of shotcrete soil penetrometer test meyco needle penetrometer test sprayed beam compres

sion test hilti nailgun pull out test and standard cylinder uniaxial compressive strength test although the sprayed beam compression test is the most reliable technique for determining in situ strength at the face the tool is difficult to handle and unlikely to be used for daily quality control clements suggested that the meyco needle penetrometer may be the more useful albeit conservative tool bernard 2008 argued that traditional empirical design methods for frs are based on term of life performance and are not necessarily applicable for early age shotcrete earlyage shotcrete fails in shear in response to applied loads from individual loose rocks as shotcrete strengthens its failure mode transitions from shear to delamination from the substrate and flexural failure thus the relationship between shear and compressive strength over the first few days of strength gain differs from that for mature shotcrete bond strength between frs and the underlying rock surface increases to a maximum of 0 2 mpa 2 bar or 29 psi at 24 hours these factors must be considered in order to establish safe reentry times dynamic testing of surface support the efnarc and rdp tests are quasi static tests where load is applied to the sample at a relatively slow rate because deep underground mines often encounter seismically active conditions where loads impact the support system via dynamic shockwaves it is important to understand the response of surface support under dynamic loading stacey and ortlepp 2007 summarized results from dynamic testing programs performed in south africa using a drop weight testing apparatus in this test the dynamicenergy impact is delivered through a pyramid of bricks to distribute load and simulate the broken rock mass surrounding deep south africans mines table 8 6 7 shows the average dynamic energy impact and displacement values at failure for different surface support systems it is important to realize that a significant amount of energy dissipates in the brick pyramid and only a fraction 50 according to human 2004 is absorbed by the support system an advantage of this test is that it examines the complete support system including interaction of the reinforcement 1 1 m bolting pattern using both stiff and yielding bolts with the surface support player et al 2008 conducted a series of dynamic tests on 1 3 1 3 m 50 50 in panels of two types of mesh galvanized wire mesh 5 6 mm diameter welded into a 100 mm 4 in square grid pattern and high strength steel wire chainlink mesh 4 mm diameter the average dynamic rupture displacement for welded wire mesh is 200 mm 8 in and for chain link mesh is 300 mm 12 in the average dynamic rupture force for welded wire mesh is 55 kn and for chainlink mesh is 185 kn the average dynamic rupture energy for welded wire mesh is 2 kj and for chain link mesh is 9 kj comparisons among different testing programs are difficult because of the different testi

ng methods shotcrete stabilization mechanisms in its original tunneling applications shotcrete was generally designed to work as a compressive arch taking advantage of the high compressive strength of concrete tunnels were generally either circular or horseshoe shaped promoting strong arching action of the shotcrete layer the surface support layer was often applied floor to floor or all around the tunnel allowing for good load transfer from the roof and wall to the floor figure 8 6 19 in mining applications however because of cost and time constraints the short term nature of excavations and the fact that the general public is not exposed the application of shotcrete differs considerably from the case for tunneling applications the floor is very rarely supported surface support sometimes does not extend all the way to the floor or when it does the layer can be very thin as a result the compressive strength of shotcrete is not really brought into play figure 8 6 20 surface support is provided through the adhesion and or shear strength between rock and shotcrete the flexural strength of the shotcrete and load transfer to the rock reinforcement stacey 2001 has identified the various support mechanisms provided by shotcrete these mechanisms can occur individually and in combination promotion of block interlock the immediate reactive support generated by the bond between the liner and the rock mass promotes block interlocking and prevents loosening of the rock mass a number of submechanisms take place in this process interlock due to bonding of shotcrete to the rock and the tensile strength of the shotcrete figure 8 6 21a shear resistance between the rock mass and the shotcrete is produced by the bonding restricting blocks from moving rotating or loosening development of shear strength on the interface between shotcrete and rock figure 8 6 21b this is the case in poorly bonded shotcrete penetration of shotcrete into joints and cracks figure 8 6 21c penetration of shotcrete into surface cracks inhibits the movement of blocks applying a thin layer of liner to contain stress induced spalling figure 8 6 21d a layer of liner applied at an early stage of fracturing inhibits the fracturing process this is particularly relevant in high stressof 0 2 mpa 2 bar or 29 psi at 24 hours these factors must be considered in order to establish safe reentry times dynamic testing of surface support the efnarc and rdp tests are quasi static tests where load is applied to the sample at a relatively slow rate because deep underground mines often encounter seismically active conditions where loads impact the support system via dynamic shockwaves it is important to understand the response of surface support under dynamic loading stacey and ortlepp 2007 summarized results from dynamic testing programs performed in south africa using a drop weight testing apparatus in this test the dynamicenergy impact i

s delivered through a pyramid of bricks to distribute load and simulate the broken rock mass surrounding deep south africans mines table 8 6 7 shows the average dynamic energy impact and displacement values at failure for different surface support systems it is important to realize that a significant amount of energy dissipates in the brick pyramid and only a fraction 50 according to human 2004 is absorbed by the support system an advantage of this test is that it examines the complete support system including interaction of the reinforcement 1 1 m bolting pattern using both stiff and yielding bolts with the surface support player et al 2008 conducted a series of dynamic tests on 1 3 1 3 m 50 50 in panels of two types of mesh galvanized wire mesh 5 6 mm diameter welded into a 100 mm 4 in square grid pattern and high strength steel wire chainlink mesh 4 mm diameter the average dynamic rupture displacement for welded wire mesh is 200 mm 8 in and for chain link mesh is 300 mm 12 in the average dynamic rupture force for welded wire mesh is 55 kn and for chainlink mesh is 185 kn the average dynamic rupture energy for welded wire mesh is 2 kj and for chain link mesh is 9 kj comparisons among different testing programs are difficult because of the different testing methods shotcrete stabilization mechanisms in its original tunneling applications shotcrete was generally designed to work as a compressive arch taking advantage of the high compressive strength of concrete tunnels were generally either circular or horseshoe shaped promoting strong arching action of the shotcrete layer the surface support layer was often applied floor to floor or all around the tunnel allowing for good load transfer from the roof and wall to the floor figure 8 6 19 in mining applications however because of cost and time constraints the short term nature of excavations and the fact that the general public is not exposed the application of shotcrete differs considerably from the case for tunneling applications the floor is very rarely supported surface support sometimes does not extend all the way to the floor or when it does the layer can be very thin as a result the compressive strength of shotcrete is not really brought into play figure 8 6 20 surface support is provided through the adhesion and or shear strength between rock and shotcrete the flexural strength of the shotcrete and load transfer to the rock reinforcement stacey 2001 has identified the various support mechanisms provided by shotcrete these mechanisms can occur individually and in combination promotion of block interlock the immediate reactive support generated by the bond between the liner and the rock mass promotes block interlocking and prevents loosening of the rock mass a number of submechanisms take place in this process interlock due to bonding of shotcrete to the rock and the tensile strength of the shotcrete figure 8 6 21

a shear resistance between the rock mass and the shotcrete is produced by the bonding restricting blocks from moving rotating or loosening development of shear strength on the interface between shotcrete and rock figure 8 6 21b this is the case in poorly bonded shotcrete penetration of shotcrete into joints and cracks figure 8 6 21c penetration of shotcrete into surface cracks inhibits the movement of blocks applying a thin layer of liner to contain stress induced spalling figure 8 6 21d a layer of liner applied at an early stage of fracturing inhibits the fracturing process this is particularly relevant in high stress situations where some loosening of the blocks would have already occurred applying a thick shotcrete layer figure 8 6 21e the internal shear strength of the shotcrete prevents larger failure applying a thin shotcrete layer figure 8 6 21f adhesion between the shotcrete layer and the rock fails locally causing the tensile or flexural strength of the shotcrete to be solicited to resist movement and loosening basket mechanism figure 8 6 21g when the rock mass behind the shotcrete is crushed and is in a yielding state the action of the shotcrete layer becomes similar to that of mesh and is required to contain the broken rock mass somewhat like a basket in this situation the toughness of the shotcrete becomes critical to resist deflection of the broken rock mass which can be exacerbated by stress changes due to stoping nearby because the tensile strength of concrete is low the liner is expected to crack early and therefore must rely heavily on reinforcement material such as mesh or fibers to contain the deflection structural arch figure 8 6 21h the arched shape solicits the compressive strength of the shotcrete to resist movement also contributing to the resisting forces is the shear strength between the shotcrete and the rock surface if only the crown of the arch is supported as is often the case in mining the support mechanism relies on load transfer between the shotcrete and the reinforcement as well as the adhesion shear strength and flexural strength of the shotcrete slab enhancement figure 8 6 21i laminated rock masses particularly when strong foliation or bedding runs parallel to one of the excavation surfaces may fail due to buckling a shotcrete layer effectively thickens the slab decreasing its slenderness and increasing its resistance to buckling extended faceplate figure 8 6 21j the application of shotcrete extends the area of influence of rock bolt faceplates it also enhances load transfer from the excavation surface to the reinforcement installing bolts after the shotcrete is also highly desirable because if the layer cracks the bolt can then pin the broken shotcrete pieces to the rock mass and prevent them from falling off and becoming a hazard the effect of an extended faceplate is lost where shotcrete is applied over mesh or frs is applied

after bolting surface protection certain rock mass may be prone to deterioration when subjected to wetting or drying shotcrete can act as a protective layer against the environment or mechanical deterioration from blast vibration equipment collision and so forth shotcrete failure mechanisms from the stabilization mechanisms just described stacey 2001 derived the corresponding failure mechanisms commonly observed in shotcrete figure 8 6 22 wedge and block loading discrete blocks or wedges defined by local discontinuities can move because of stress and or gravity and load the shotcrete layer at specific locations figure 8 6 22a as noted previously if the shear strength and adhesion between the rock mass and the shotcrete holds failure is likely to occur by shearing of the shotcrete layer if the shotcrete layer debonds flexural failure of the shotcrete is the likely mechanism distributed surface loading distributed loading of the surface can be due to a number of rock mass yielding mechanisms such as gravity acting on failed rock squeezing ground under high stress or dynamic loading from seismic events this mechanism is related to the basket mechanism figure 8 6 22b the surface support must work with the rock reinforcement as a system and unless a good arch is formed by the liner failure is likely to be by flexural bending stress induced loading if a strong bond is achieved between the shotcrete liner and the rock mass the two deform as a unit because concrete is brittle a number of possible stress conditions can develop within the shotcrete layer resulting in different modes of failure including the following shear failure figure 8 6 22c bending failure in the roof figure 8 6 22d and buckling failure in the walls as a result of floor heaving figure 8 6 22e complex failure modes involving combinations of the above are also possible in rock bursting conditions a common failure mechanism for shotcrete is slab ejection figure 8 6 22f if a shotcrete shell is not properly drained in areas where groundwater is flowing water pressure can build up and cause flexural failure in high horizontal stress conditions where significant floor heaving and wall convergence are observed the shotcrete shell will bend with the inward wall displacement resulting in the crushing of the shotcrete at the corners of the excavation figure 35 22g design of shotcrete liners load and failure mechanisms involving liner support can obviously be quite complex shotcrete liners can be designed either empirically or deterministically empirical design approaches a reasonable starting point to an empirical approach is to consider the common practice of applying reinforced shotcrete of thickness 50 mm 2 in or 100 mm 4 in if the excavation span is large in short life mining excavations hoek 2007 lists recommended shotcrete applications in underground mining for different rock mass conditions several mines rely on rock

mass classification systems to define rock mass quality and select what they consider appropriate shotcrete support to complement their reinforcement strategy one such system the q system proposed by barton et al 1974 makes specific recommendations for the use of shotcrete and mesh reinforced shotcrete in combination with rock bolts grimstad and barton 1993 revised the q system to encompass the use of frs which was an unknown product when the system was first published in 1974 see figure 8 8 8 in chapter 8 8 when frs is used the spacing of the rock reinforcement rock bolts can be expanded by 20 40 compared to the case for bolting only grimstad and barton 1993 proposed an approximate relationship between frs thickness as a function of rock mass quality q and excavation size figure 8 6 23 melbye et al 2001 proposed that when high performance steel fiber reinforced shotcrete is used as a wet mix the required thickness proposed by grimstad and barton can be further reduced figure 8 6 24 grant et al 2001 argued that the recommendations of grimstad and barton 1993 were inadequate because they did not consider the impact of the toughness of frs papworth 2002 expanded this work by including reference to the rdp test figure 8 6 25 provides further guidelines on the recommended frs toughness for different rock mass qualities as defined by the q system in areas of anticipated significant deformation seismicity or potential strain burst a minimum energy absorption capacity of 400 j should be used based on rdp40mm in extreme cases this should be increased to 540 j papworth further suggests that wherever possible bolting should be done through the frs deterministic design approaches fernandez delgado et al 1981 proposed a deterministic model to calculate the shotcrete thickness t required to support the dead weight of a wedge based on arching theory and the compressive strength of shotcrete for the rock reinforcement scenario shown in figure 8 6 26 barret and mccreath 1995 proposed deterministic design models to analyze the worst case scenario of discrete blocks that could develop in between the rock reinforcement the models consider four failure mechanisms adhesive direct shear flexural and punching shear figure 8 6 27 model calculations are shown in table 8 6 8 shotcrete support for rock wedges software is available to conduct kinematic and limit equilibrium wedge analysis for underground excavations an example of this type of software is unwedge rocscience which investigates the influence of both rock reinforcement bolts and surface support such as shotcrete a basic assumption is that shotcrete fails in direct shear a structurally defined wedge shears through an area of shotcrete that is given by the product of the perimeter of the exposed face and the thickness of the shotcrete the shotcrete acts in a direction normal to the excavation wall that it supports numerical modeling app

roaches several numerical models can model shotcrete either individually or as part of a rock bolt support system these models are more often used to investigate the impact of different shotcrete characteristics and properties on the stability of excavations rather than for design purposes kuchta et al 2004 used a distinct element approach to construct a simple drift model containing a triangular wedge in the crown the model evaluates the relationship between shotcrete thickness and interface strength and demonstrates that tensile bond strength is more important than cohesive bond strength in determining interface strength malmgren and nordlund 2008 performed a comprehensive numerical study on the interaction of shotcrete and rock bolts related to conditions at the lkab underground mine in kiruna northern sweden their models demonstrate that rock strength and surface roughness have a significant impact on the number of failures at the rock shotcrete interface and in the shotcrete lining they also show that shotcrete lining is sensitive to small amplitudes of surface roughness the implication is that shotcrete performance depends on the method of excavation drilling and blasting and the quality of ground preparation in all cases high interface strength is favorable because the performance of thick shotcrete lining depends on bond strength there are limited benefits to using thicker shotcrete lining shotcrete performance the performance of ground support systems in underground mines is generally intimately related to the quality of their installation and shotcrete is no exception some of the guidelines relevant to quality control of shotcrete are briefly discussed in the following sections quality control shotcrete performance in underground mines is influenced by the implementation of appropriate quality control procedures concerning design materials application equipment craftsmanship and installation techniques spearing 2001 identifies the following requirements for a quality control program it must be reliable meaningful timely simple and relatively inexpensive for shotcrete designs derived either empirically or analytically quality control ensures that shotcrete is applied as designed other surface support techniques a number of specialized techniques have been applied as surface support in underground metal mines albeit to a lesser extent than mesh straps and shotcrete three such techniques are briefly discussed thin spray on liners shotcrete pillars and reinforced shotcrete arches thin spray on liners thin spray on liners tsls are multicomponent polymer materials that can be applied to the rock mass surface as a sealant or as surface support after more than 20 years of research development and trials in mines the use of tsls in underground metal mines is still limited to special applications rather than to general application as a systematic surface support such as mesh and shotcrete ne

vertheless their popularity in certain coal mines and in some less mechanized south african mines has been growing in recent years the evolution of tsls until 2004 is covered in potvin et al 2004 which includes product reviews testing and quality control procedures and case studies a main advantage of tsls is their rapid and easy application since only a thin layer 5 mm is required the equipment used for application is small and versatile a variety of tsl products exist with different chemical components and properties they can be divided into two categories those that contain isocyanates and those that do not the first category includes polyurethanes and polyresins the second category includes water cement or latex based methacrylates and acrylates with regard to structural properties tsls that contain isocyanates are generally superior to those that do not unfortunately however isocyanate products have serious toxicity issues because they contain methyl disphenyl isocyanate mdi which in part of the population can induce a severe allergic reaction and even death therefore rigorous health and safety procedures must be followed if these products are to be used tsls have been used successfully for specialized applications including sealants in south african mines and over shotcrete in canadian mines to further safeguard against cracking in critical excavations a review of the literature indicates that tsl performance assessments are not consistent in particular the critical properties and the laboratory and field tests required to estimate these properties are still the object of discussion among scientists and practitioners the main suppliers of tsl products currently include minova tekflex basf masterseal sa mining engineering supplies tunnel guard 3m 3m mining liner and spray on plastics ltd rockweb most of these products are still under study and field trial and no reports yet exist about tsls being systematically applied as surface support in mines therefore the technology should be deemed not yet mature and still under development shotcrete pillars used mainly in canadian underground metal mines are a specialized form of surface support that does not cover the entire rock mass surface rather it is external to the rock mass and stabilizes the roof of an excavation use of this method of support depends on factors such as excavation access potential rock mass failure mechanism support capacity required and economics pillar construction is very simple on solid rock a pillar is built as a column of shotcrete with a diameter including final cover of shotcrete in the range of 1 3 to 1 5 m 4 25 to 5 ft figure 8 6 29 casing when used is made of a 9 mesh cylinder of slightly smaller diameter contacts at the floor and back are slightly expanded to give it an hourglass shape methods for calculating critical design parameters such as pillar dimension and pillar reinforcemen

t are presented in hadjigeorgiou et al 1999 figure 8 6 30 shows a graphic tool for estimating the capacity of mesh reinforced shotcrete pillars as a function of pillar diameter and shotcrete compressive strength reinforced shotcrete arches reinforced shotcrete arches provide stiff high capacity surface support they can be installed quickly and efficiently in an underground mine gaudreau et al 2004 examined shotcrete arches at the brunswick canada mining and smelting operation including their method of construction they are assembled by juxtaposing prefabricated steel frames made of 3 reinforcing bars and 0 mesh straps figure 8 6 31 reinforcing frames are secured to the rock mass surface by means of friction bolts or grouted bars and plates to form a complete arch from floor to floor the arch is then filled with shotcrete to a thickness of 50 mm 2 in starting from the base until the entire frame is covered double or triple arches can be built side by side to increase capacity and improve coverage of the supported area when necessary such as for large spans or very difficult ground conditions table 8 6 10 lists the general rules used at the brunswick mining and smelting operation to select single double or triple arches according to the span of the excavation when multiple arches are used they are typically installed every 3 m 10 ft but this distance can vary when mining in extremely poor rock mass or when rehabilitating collapsed areas arches can also be installed in combination with shotcrete pillars offering versatile solutions to difficult ground conditions summary reinforcement and surface support interact to provide an integrated ground support system their roles are dictated by the mode of interaction with the rock mass whereas reinforcement elements preserve the integrity of the rock mass surface support links reinforcement elements together to resist surface deformation or to contain rock fragments from falling or ejecting the main surface support elements used in the metal mining industry are welded wire mesh shotcrete and to a lesser extent steel straps important properties of surface support include stiffness load capacity and deformation capacity welded wire mesh is soft with high displacement and relatively low load bearing capacity its support action generally starts only when the rock mass has failed shotcrete is stiff with low deformation capacity and relatively low loadbearing capacity in mining applications because the concrete is often not loaded in compression its support action begins as soon as the rock mass starts to deform straps do not cover the entire surface and are often used over mesh or in special applications innovative support techniques such as shotcrete pillars and reinforced shotcrete arches provide more ground control options thin spray on liners still under development are used primarily for special applications in metal mines although they are increa

singly used in coal mines and in deep underground south african mines ground control is the science that studies the behavior of the rock mass in transition from one state of equilibrium to another it provides a basis for the design of support systems to prevent or control the collapse or failure of roof floor and ribs both safely and economically peng 1978 around the world many mined minerals occur as soft rocks themselves or are found within a rock mass comprised largely of soft rocks the mining of these resources presents many unique problems significant among these is the need to provide supports that are both temporary and permanent their installation being part of the overall mining system this chapter deals with this key issue of support by examining the types of mining system that have become widely used around the world to extract these important minerals it covers changes in stress distribution and rock displacement brought about by mineral extraction and the types of support system that have been used successfully in this complex environment it concludes by examining planning and design methodologies for support in soft rocks in the interests of brevity and efficiency this chapter concentrates on the working of commonly found sedimentary deposits such as coal but with occasional reference to the working of evaporite deposits what is soft rock rock is an engineering material having a unique set of engineering parameters these parameters most commonly include not only uniaxial compressive strength but also tensile strength stiffness young s modulus the ratio of lateral to vertical strain under load poisson s ratio and behavior when confined which is defined by the angle of friction and its cohesion however these properties are not the only factors that define the behavior of this material when subjected to stress changes brought about by mining material properties measured from small samples in a laboratory give only guidance to a wider understanding of that rock type s behavior within the rock mass the pattern and frequency of jointing within a rock mass together with the presence of geological anomalies such as faulting folding and weathering greatly affect the way a rock type might behave in a certain mining application these factors have been quantified within rock mass classification systems the most significant of which are the rock mass rating rmr system developed by bieniawski over many years and the norwegian q system developed by barton the presence of water has a particularly large effect on rock stability as many soft rock types are negatively affected by exposure to water another element influencing mining support in soft rock conditions is time dependency excavations in soft rocks become subject to closure with time due to characteristics of the rock and of the ground mass the rock types themselves exhibit creep properties that is the propensity for strain to increase under a co

nstant load this is most notable in rocksalt deposits additionally variations in the stress fields within the ground mass and subsequent development of fracturing due to the mining process also result in deformation of the excavation the support applied to excavations in soft rock must take this movement into account the general term soft rock is often used to describe a group of rocks with the following specific properties but it is also a comparative term used to contrast with the other widely used term hard rock soft rocks have uniaxial compressive strengths determined by laboratory testing in the range 50 to 0 6 mpa below which the material is termed very stiff soil this range can be further subdivided as follows moderately weak rock 10 50 mpa weak rock 1 5 10 mpa very weak rock 0 6 1 5 mpa in the rmr system within the soft rock range rocks with an rmr in the range 20 40 are described as poor rock and those with an rmr of 20 are described as very poor rock from the point of view of support the yield characteristic of a soft rock is more nearly plastic as opposed to brittle the latter tends to be more characteristic of hard rock the terms soft rock and hard rock are often used to determine the difference between two groups of mines and their mining methods soft rock mines include all coal mines plus potentially mines using typical coal mining methods such as longwall and room and pillar these might include as examples mines of gypsum rock salt limestone and potash the strata are sedimentary in origin hard rock mines include those using nearly all the metalliferous mining methods such as stoping in its many forms the rock types are almost exclusively igneous or metamorphic these are very general classifications and although there are exceptions the general rule normally applies many examples exist around the world of coal being worked in a rock environment in which a substantial proportion of the rock types exceeds the arbitrary values listed above however the problems of mining support in soft rock are significant and it is the solutions to these that form the basis of this chapter theoretical background this section examines the stress changes around an excavation caused by mining and the influence of these changes on roadway closure the study of stress change and strata displacement around soft rock excavations is very mature much of the work having commenced up to 80 years ago and being developed and refined in the period after the second world war and up to the 1990s as a result although the references quoted and images used may seem rather elderly to the younger mining engineer their relevance has not diminished with time nor have the findings been superseded stress and displacement around an excavation in soft rocks found in coal measures or evaporite deposits stresses as redistributed into the rock mass often exceed the elastic bounds of each rock type in these circumstances

support plays a significant role in the stability of the excavation soft rock mechanics is often called strata control in the longwall method of mining there is substantial controlled rock mass movement from the excavation level through to the ground surface in built yield is a key design element of many support types used in the soft rock domain this process of controlling rock mass movement is particularly evident when considering choices of support at a longwall face and in the roadways serving the longwall early researchers in the united kingdom characterized the redistribution of vertical stresses around a coal mining excavation as a pressure arch national coal board 1950 although this theory was later modified the fundamental redistribution of these stresses around an excavation to form high stress concentrations known as abutments around the perimeter of the excavation is widely recognized an understanding of the nature and role of the components of horizontal stress is also important although horizontal stress did not really become widely understood in the european coal mining community until the major move to rock bolted support at the end of the 1980s this understanding was based primarily on the work of gale 1991 in australia in the worst cases high stress concentrations at the sides or ribs of an opening caused rib side failures resulting in effective widening of the excavation horizontal stresses that become concentrated in roof and floor measures cause buckling of those rocks above and below the excavation as evidenced by roof buckling and floor heave figures 8 7 1 and 8 7 2 the degree of closure or rock failure depends on the stress values and the strength of the rocks horizontal stresses are often characterized by directionality the component of stress in one direction having a significantly higher value than the component in the direction at right angles this can dictate the choice of mine layout as there are advantages to driving the most important roadways parallel to the maximum component direction to minimize closure also in some shallow deposits the magnitude of the maximum horizontal stress can be greater than thatt of the vertical stress which depends largely on depth the role of the support member in soft rock excavation is to accommodate strata displacement in the roof and sides of the excavation in a controlled manner rather than to prevent it entirely the support action of the rock bolt is an exception to this rule support of the floor has always been viewed as technically difficult because any system adopted will interfere with mining activities and infrastructure within the excavation this support is also difficult to replace or renew should failure occur support placed to the sides and roof of a working area or access roadway is normally easier to supplement or replace interaction interaction is a mining phenomenon in which the working layout at one horizon influences the st

ability of workings at another horizon in the mine this phenomenon is normally experienced only in the mining of stratified deposits wherework is conducted at more than one horizon in the deposit it is a common feature of coal mining where several seams are worked in the vertical succession planning for interaction has been a feature of coal mine design in the united kingdom and europe for many years it is increasingly recognized as a constraint on the development and working of coal reserves in the united states south africa and australia where second seam working is now practiced more widely interaction can be subdivided in a number of ways it has been described as being split into static project and dynamic project activities auriol 1972 a static project is one in which an existing working passes into a zone in which the field stress has been modified by previous workings often long since abandoned these can be old workings above or below the present working level a dynamic project is one in which existing excavations working areas or main roadways can be influenced in some way by current working activity at another horizon as a result of stress redistribution ground movement or both interaction can take a number of forms it can be solely a redistribution or modification of the stress field being encountered it can also be induced fracturing caused by previous workings in addition in a dynamic project if the affecting workings are below a horizon it can be ground displacement caused by the onset of subsidence ground displacement can also lead to induced ground fracturing in the form of tension cracks formed around excavation edges and possible mobilization of fault planes allowing for pathways to be formed for water or gas flow into existing workings in early observations of interaction researchers noted its effects and attempted to ascribe theories to what they had experienced underground the empirical studies of walker and associates in the north east england coalfield of the early 1950s were typical of this as espoused in their pressure arch theory mentioned previously national coal board 1950 1951 observations of damage to mining excavations suggested that the pressure arch extends both above and below the working horizon and any subsequent working in another seam that intersects the pressure arch experiences difficult mining conditions national coal board 1953 1954 the extent of the pressure arch above and below workings was given as twice its width as an example for workings at a depth of 610 m the pressure arch is said to extend 220 m above and below the seam the magnitude of the stress levels in the pressure arch legs is related to depth varying from 1 5 times the cover load at shallow depths and stress concentrations to over 3 times the cover load at depths of 600 m the maximum extent of the field data obtained by the investigators subsequent researchers most notably whittaker took thi

s theme further publishing a set of values showing increased stress at levels above a single rib edge he states the following the mirror image of this pressure profile is valid for the state of stress underneath the solid rib side except that the magnitudes of the pressures are slightly increased owing to the greater depth whittaker 1972 king and whittaker 1970 published a diagram that demonstrates the development of stress contours below a remnant pillar figure 8 7 3 the large body of subsequent published work especially that appearing in the yearly proceedings of the annual conference on ground control in mining which is hosted by the university of west virginia united states and the u s bureau of mines and its successor bodies and is regarded by many as the premier world forum on soft rock mechanics research reviews the effects of interaction and its influence on working layouts and excavation stability this work is dominated by reviews of experience in u s mines with occasional australian south african and european examples other research associated with multiseam mining interaction can also be found in the final reports of some of the research projects funded by the european coal and steel community up to 1992 thereafter the research fund for coal and steel rfcs stress prediction is important but still must be related to its effect on basic roadway stability in the affected zone that is one can mine through a high stress zone and stability may not be affected or the reverse might apply until the late 1980s in u k coal mines roadways were normally formed to an arch profile and supported by steel supports roof closure experienced in advancing gate roads of longwall coal faces frequently amounted to more than 50 of the extracted seam section floor lift was also commonplace closure therefore was expected and a meter of convergence was not uncommon an additional increment of closure of a few centimeters due to interaction stresses was often not noticeable unless measurements were taken when rock bolts became the principal support in rectangular section retreat roadways and measurement of closure and adjacent ground movement became routine the effects of interaction began to be recognized with ground movement action levels at which additional support is to be set now quoted in millimeters any additional roadway closure due to interaction effects is identified and significant the influence of interaction is therefore related to the fundamental stability of the affected roadway mining methods soft rock mining methods can be subdivided into longwall and derivatives of rock and pillar longwall mining throughout the world the longwall method of mining produces the highest levels of bulk tonnage of mineral particularly but not exclusively in coal mining coal was worked for many years by mining systems that relied on manual effort for coal extraction and transport these systems included the extraction op

eration elements that form the origins of today s longwall face from the invention of the armored flexible conveyor power loader and powered self advancing support in the late 1940s and early 1950s the system of longwall operations that we know today emerged single longwall faces between 100 and 150 m long served by two gate roads in the advancing mode became the european standard the gate roads were formed behind the face line at each extremity of the longwall usually by the removal of seam roof to give additional height for access and ventilation working height at the coal face was limited not only by seam thickness but also in thicker seams by support designs available for use on the longwall face support of the gate roads was provided by steel standing supports usually in an arch configuration on one or both rib sides where coal had been extracted packs were built from stone timber or other imported materials fundamentally little has changed in longwall mining except for the equipment used higher powered more robust flexible conveyors now allow face lengths of more than 300 m more powerful power loaders have increased production by virtue of their speed the width of web they cut and the height that they can extract from the seam in a single pass powered roof supports now have yield capacities of 1 000 t per unit and are capable of supporting working heights of 5 m at the face additionally the major underground coal producers of the world have all chosen to practice retreat rather than advance longwall mining in a retreat system gate roadways are driven to the extremity of the reserve prior to the commencement of coal extraction distances of 5 km are not uncommon longwall faces are then formed in the seam to work back between them after which they are abandoned in the longwall waste roadways for retreat faces can be single drivages formed at or near seam height with a rectangular cross section or multiples of two three or four parallel roadways connected at frequent intervals by crosscuts single gate roadways are commonplace in the deeper workings of europe figure 8 7 4 whereas multiple gate roadways serving the longwall are commonplace and legally required in mines where the coal seams are substantially shallower as in the united states and australia figure 8 7 5 the versions of longwall mining described above were appropriate for coal seams 5 m thick laid down as near horizontal deposits systems have been developed to work exceptionally thick seams by the longwall method using derivatives of the sublevel caving idea in which caving waste which is comprised largely of coal is accessed through the rear shield of the powered support and loaded onto the face conveyor many seams around the world most notably in northern spain are highly inclined versions of the longwall method are used with individual gate roadways maintained on the strike of the seam with the production face at or near

full dip figure 8 7 6 in the united kingdom there was always considered to be a limiting seam angle of inclination in the range 45 50 beyond which relatively standard longwall techniques and equipment could not be used above this angle specialist systems and equipment were required as production from relatively thick coal seams that lie in near horizontal deposits can be obtained in substantial quantities and at very low cost by longwall methods extraction of coal from thin seams or from steeply inclined or geologically disturbed deposits that require special techniques has waned in importance thus this chapter concentrates on standard support techniques required for higher performing longwalls room and pillar mining the alternative to longwall soft rock mining methods comes under the generic heading room and pillar r p r p methods require the driving of multiple parallel roadways in the mineral deposit joined by crosscuts to form a matrix of unworked pillars of mineral that are stable enough to provide overall long term support for the mining section figure 8 7 7 this is fundamentally an advancing mining system when mining operations have extended to the planned or natural boundary the choice remains to either abandon the section or operate a secondary mining system that removes some or all of the pillars upon retreat secondary mining requires specialized support systems in which waste is formed by extraction of the support pillars the extent of this waste must be controlled by temporary supports that separate the remaining pillars and accessible roadways from those where caving is occurring in recent times mobile powered roof supports have been used for this purpose with great success many alternative r p versions exist some involve a herringbone pattern layout some have pillars that are deliberately smaller than would normally be regarded as large enough for long term stability and some have short headings driven into the flank of the panel to maximize extraction support requirements for some of these systems are covered in this chapter but individual systems are too many and varied to be covered here in detail r p methods rarely match the production capacity of longwall methods but they do offer a low cost alternative where capital investment is limited or some restrictions exist on the choice of mining method because of the need to preserve surface features from the effects of subsidence r p methods are also used where the roof is so competent that it would not be expected to cave in a regular and controlled manner if longwall were used the major restriction on the use of r p methods is the depth at which they can be efficiently and safely used it is unusual to find r p methods used in coal mining at depths greater than 500 m from the surface below this depth the pillar sizes required to support the district become so large that percentage extraction becomes too low and distances between each advancin

g heading become too high for efficient operation of the mobile plant normally found in these systems additionally the cost of support required to maintain safe conditions in all the roadways mined can become significant interestingly some evaporite deposits most notably potash are being worked at much greater depths 1 000 1 300 m by r p versions in which pillar sizes are small enough to yield with time figure 8 7 8 choosing a type of support a brief introduction the support that is selected for mining excavation has a number of roles to satisfy it must apply constraint to the yield zone of failed rock immediately next to the excavation to assist in the natural strengthening of that zone and enable the excavation to remain stable for its designed life whether 3 months or 50 years it must provide sufficient working space for mining equipment and operations it must prevent additional small pieces of loose material from the surface of the excavation from falling and striking people and machinery finally it must provide people who travel through or work in the excavation with the confidence that they can do so safely a wide range of support options are available for softrock mining perhaps the only option that rarely arises is that of no support being needed examples of available support types are discussed in the following paragraphs rock bolts reinforce the rock mass giving it additional strength to withstand imposed stress concentrations they are designed primarily to resist rather than accommodate strata yield rock bolt systems are highly flexible patterns being altered to react to localized changes in support needs they are relatively cheap and easy to transport but require skill and supervision to install there is also some opportunity for numerical design rock bolts are frequently used in conjunction with steel mesh panels or lengths of geogrid that provide protection in the working area from spalling roof and sides of the excavation steel standing supports steel ribs are commonplace in softrock excavations they are erected from sections bolted together on site and can be easily adapted to variations in roadway size or drivage direction although in earlier times they could be found in district development and gate roads in most european coal mines the wider application of rock bolts has replaced them in a number of shorter term roadways steel standing supports can incorporate yield mechanisms to accommodate rock movement around an excavation and are normally constructed from steel grades that fail in a ductile rather than brittle manner these support types are relatively inexpensive and operate most effectively with backfill or in smoothly profiled excavations that give multiple points of contact with the rock design for these support types tends to be empirically based injection of resins and grouts into fracture planes to improve strata strength has achieved much success in specific locations such as t

he area around the end of the retreat longwall face the success of the technique depends on the presence of existing rock fractures thus it can be used not to prevent initial development of fractures but only to prevent fractured rock from becoming more fractured and threatening the stability of the mining excavation it is normally used in conjunction with rock bolt systems and frequently requires specialist operators and equipment to achieve successful results concerns exist about the sensitizing properties and flammability of many of the more successful injection products especially those based on polyurethane targeting of injection in the required areas is difficult the cost of the chemicals and specialist assistance makes injection relatively expensive hydraulics revolutionized the support of longwall face roofs a longwall requires temporary moveable and yielding support and supports based on hydraulic cylinders meet all of these criteria single hydraulic props were replaced in a 50 year evolutionary process by powered roof supports that support the longwall working area advance the face conveyor and then advance themselves as extraction continues versions of powered supports have been used as mobile temporary supports for extracting pillars from r p systems sprayed concrete shotcrete can be applied as an initial temporary support measure normally in a major drivage to provide instant support until secondary more permanent systems are erected it plays a key role in large excavations such as tunnels large roadway junctions or shaft insets where fastacting good quality temporary support is needed this system requires skilled operators and specialist equipment and products a potential drawback is that there is little scope for yield when ground eventually moves and the fractured shotcrete liner itself becomes a safety hazard concrete or cast steel segments are the ultimate in highstrength permanent lining and are used mainly in shafts major long life underground drivages and places where water inflow might threaten the mine such as when driving through major faulting or driving shafts through water bearing rock these linings are expensive they are specially designed for a specific excavation and normally require backfill or backgrouting to prevent point contact loads from destroying their integrity they are inflexible in that they are difficult to install in other than straight line excavations and it is difficult to vary the dimensions of the opening they require major logistical backup to bring them to where they are to be erected and special erection arrangements must be provided packs constructed from stone timber concrete blocks or pumped cementitious material have been used in many parts of a mine as supplementary support stone packs were originally a key part of longwall support when built as strips in the waste before the advent of powered supports they were also used as additional roadside sup

port in roadways of advancing longwall faces in modern times packs are frequently used in places such as retreat face ends where significant additional support is required to retain access ahead of or ventilation flow behind the face line their key properties are their flexibility and their significant support load carrying capability but they are labor intensive to build and require import of materials the following sections examine in more detail some of these key support types rock bolts rock bolts are widely used in soft rock mining both as principal support and as supplementary additional support or temporary rock reinforcement their primary function is to reinforce near seam rock so that it can experience loading without excessive deformation and hence deterioration to the extent that the mine opening is damaged peng 1978 describes three main methods of support action 1 suspension of a soft rock beam from a more competent upper bed 2 formation of a composite beam by rock reinforcement 3 increase in the frictional resistance of jointed rock across failure planes to help promote natural arching rock bolts were widely and successfully used in coal mines in the united states from the mid 1940s importantly they complemented the normal r p system of mining they were also appropriate to strata strengths and stress conditions in u s mines rock bolts were initially introduced in the united kingdom in the 1960s as mechanical anchorages for which use they ultimately failed they were reintroduced in the mid1980s because competitive pressures on u k coal producers were driving the industry to rely on retreat longwall mining to deliver the required volumes of output at prices that the market could stand the industry determined that the australian experience with rock bolted support in coal mines was appropriate for u k conditions these conditions dictated that rock bolt systems must resist rock movement whereas existing standing supports accommodated movement there is very little scope for yield in a rock bolt system and their success was judged on the degree of stiffness in the system only rock bolts that are full column anchored with resin have been used successfully in u k coal mines their performance as a support in high stress soft rock conditions is superior to that of alternative forms of rock anchorage much debate has ensued as to whether tensioning of a full columnanchored rock bolt after installation contributes to its ultimate capability to support the rock strata research using numerical modeling saleh et al 2004 has shown that the effectiveness of a pre tension load varies with both rock mass quality and in situ stress level the findings can be summarized as follows the compression zone generated by tensioning a fullcolumn anchored bolt is very limited in extent pre tensioning offers some advantages in a closely spaced bolting pattern in a weak jointed rock mass rmr 30 pre tensioning also

offers advantages in a low in situ stress field mixed systems of support involving standing steel supports supplemented by rock bolts have been used in the united kingdom the support actions of each system differ and therefore do not necessarily complement one another because of these differences the total support generated is not simply the sum of the contributions of each of the two systems in the early stages of rock movement around a roadway most resistance is provided by the rock bolts when the roof has moved sufficiently to load the standing support the bolts have often ceased to be effective only a mixed system incorporating adequate support density to accommodate each of the components separately can be expected to deal with a very heavily loaded roadway other rock reinforcement methods used in u k coal mines have included the following reinforcement of geological fault zones encountered by the longwall that have led to areas of broken roof and slabbing of the coal face by means of 2 m long hardwood dowels full column anchored into the affected area this method provides some cohesion to the rock mass so that mining can continue however its use has largely ceased as powered support designs have improved injection using resins and grouts this method has found considerable success as a means of artificially bonding strata in the immediate vicinity of the face end to enable control of strata movement during mining a bonding agent based on resin or grout is pumped under pressure into boreholes that are drilled into the roof and rib sides of the retreat roadway this method is often used to best effect in conjunction with steel or cuttable rock bolts it has been extended in extreme cases to the recovery of falls on longwall faces support of not only the roadway s roof but also its rib sides in high stress soft rock conditions such support is important because high vertical stresses can create a fractured zone deep into the rib side on the longwallface side of the retreat gate road support is achieved by means of cuttable bolts made of glass reinforced plastic full column anchored into the rib side during roadway formation the axial strength of these bolts is comparable to that of steel bolts used in the roof however they can be easily cut by the shearer when it reaches the roadway cable bolts and flexible bolts the length of rock bolts used in roadways is normally limited by the height of the driven roadway unless some form of coupling is used to allow installation their use is appropriate when a combination of high strata stresses plus weak rock surrounding the excavation exists and tendons longer than 2 5 m are required these longer tendons have been used to supplement steel rock bolts to provide principal support in heavily loaded roadways cable bolts can provide support at depths of 8 m full column anchorage can be provided by pumping grout to the back of the hole after installation of th

e tendon the flexibility of the cable allows the tendon to be bent during installation flexible bolts tend to be shorter steel tendons 4 m long based on steel rope they can be spun into resin capsules inserted into the hole injected after installation or a combination of both steel standing supports development of steel standing supports has been led by the coal industries of the united kingdom and germany however since the 1940s both countries have differed radically in the choice of standing support type used in roadways in the united kingdom although advancing longwall mining was the predominant mining method it was normal to support gate roadways with a two piece or three piece d shaped steel rib known as an arch the german industry also used a form of steel rib but whereas the united kingdom used rolled steel joists rsjs of h shaped cross section figure 8 7 9 based on those specified for structural steelwork in civil engineering applications the german industry favored joists of v shaped cross section with an inbuilt yielding mechanism at the joints this latter has become known on both sides of the north sea as the t h yielding arch after the original supplier toussaint heintzmann figure 8 7 10 after the move to retreat mining began in earnest in the united kingdom in the mid 1980s and in seam rectangularsection roadways became an operational necessity two major changes were effected one was the introduction of rock bolts as principal support in roadways the other altogether less obvious and less heralded was the move to flattop or cambered top steel arches figure 8 7 11 despite the increased use of rock bolting since the 1990s many gate roads continue to be supported solely with arches or with a combination of rock bolts and arches a so called mixed support system in the united kingdom arches also continue as the system of choice in many main trunk roadways all cross measure drivages all drivages through goaf known as scours all roadways formed by blasting and all advancing face gate roads the british colliery arch the colliery arch was based on the rsj specified in british standard bs 4 1980 h section joists or i section joists as they are known in the civil engineering world were designed to have the highest possible moment of inertia in the plane of bending this requires that as much of the steel in the section as possible be away from the neutral axis in civil engineering applications there is normally only one plane of loading thus most of the steel is in the flanges in long thin webs as far as possible from the neutral axis in mining applications loading usually occurs in more than one plane and is often unevenly spread and of a magnitude sufficient to cause the joist to yield therefore mining section rollings always have most of the steel concentrated in the web and tend to have overall dimensions for flanges and webs set at values near one another e g 127 114

mm 152 127 mm a significant property of any standing support steel is its capability to fail in a totally ductile manner with no sign of brittle fracture therefore most colliery arches whether d shaped or flat topped are constructed of grade fe 430 b steel per bs en 10025 arch supports are made by cold bending steel rollings into roadway supports this involves forming fishplate and strut holes in the web and sawing the arch into appropriate lengths for ease of handling and assembly supports are normally supplied into the mine in arch packs that is collections of arch segments packed together with fishplates bolts struts stilts if used lagging and where appropriate rock bolts and resin that can support a given cycle of advance in the drivage of a roadway arch joints arches are formed from two three or even four sections of joist generally connected by means of fishplated joints however these joints can become points of weakness affecting the overall strength and operational performance of the support unit in the early days of d section arch supports all fishplates were of an in web rolled steel design the wrap around fishplate introduced in the united kingdom in the mid 1970s gave additional strength to the fishplated joint especially in the plane parallel to the roadway axis it also satisfied the requirement for a fishplate that can be loosely assembled enabling an arch to be erected in one piece a technique known as whole arch setting figure 8 7 12 in the early 1980s the pressed steel in web fishplate was introduced stronger than other existing fishplates this one was also lighter and cheaper a combination of qualities achieved by use of alloy steel containing a percentage of boron joint strength in both planes was greater than that of the original sections pressed steel plates were also used to form the chevron joints of camber top and flat top supports other arch accessories arch supports are separated by spacers known as struts the number of struts used between each setting normally depends on the length of the arch perimeter struts are either bolted through the joist web or clamped onto the joist the strongest and most expensive struts are tubular fitted with flanges at either end that are bolted through the joist web laboratory tests show that a full set of struts correctly assembled and fitted to arches can contribute significantly to support system strength a chief drawback for roadway supports mentioned previously for h section joists is the lack of a yield mechanism in the support system as opposed to the yield of the steel itself one way to overcome this is to use stilts in theory when fitted to the base of the arch legs stilts allow a degree of yielding of the overall system without threatening the integrity of the arch joist or its joints many stilt designs have been made however all were expensive yet incapable of providing consistent yield under uncontrolled field con

ditions thus the use of stilts eventually ceased although the use of stilts ceased the problem that they were designed to solve did not disappear the yield mechanism in arches today is provided by packing material between arch and strata or by arch legs that penetrate the floor the latter mechanism is thought to contribute to floor lift in roadways roadway support system strength with so many variables present in the roadway support system including the most significant external variable the loading condition to which each arch in a system is subjected determination of the strength of a single support or a number of adjacent supports is based more on engineering experience than on science all conclusions about product or system strength tend to be comparative between 1986 and 1991 the european coal and steel community funded a research project titled testing roadway support systems project objectives were to provide basic data on the fundamental strengths and characteristics of a range of roadway supports and ancillary equipment project conclusions are given below british coal corporation 1991 the most severe loading condition for a single arch is a single point load as the number of loading points increases and side restraint is offered the capacity of an arch increases accordingly the most favorable condition is a uniformly distributed load around the support support capacity can be increased by reducing support base width reducing support centers and increasing section size for fishplated joints increases in joint strength up to 80 of the section strength increase overall support strength further increases in joint strength do not further increase overall support strength struts add to the ultimate strength of a support system stronger struts have the most favorable effects these conclusions although already fairly well known to mining engineers were quantified for a range of support types and loading conditions german practice the t h section arch joist was first introduced in germany in the early 1930s the principle behind arches made from this section is to allow the portions of the arch to slide relative to one another at the joints making the arch effectively yield under load without damage to the section itself the joints are formed by overlaps between the arch legs and the crown and are secured by clamps resistance to yield can be varied by adjusting the torque on the clamp bolts the t h section also enjoys the advantage that the strength of the section is similar in both major axes visitors to german mining operations would be hardpressed to see a roadway support other than the t h section german mines are laid out as longwall systems gate roads frequently serve two longwalls and are thus subject to considerable closure they are therefore formed initially much larger than are u k roadways with the expectation that some closure will take place the t h section allows major closure w

ithout the arch losing its support integrity or suffering excessive damage the roadway is often backfilled after formation to control the loading condition on the arches t h section arches have some drawbacks they are set with the base of the v inward toward the roadway this causes difficulties in the provision of struts and devices to hang cables and hoses however their chief drawback is that the yielding process because it relies on friction is difficult to control the definitive english language work on roadway support in germany is a book published in the united kingdom by fritz spruth 1960 spruth identified the importance of the joint friction clamp that allows yield and stated that clamps should do the following 1 yield at a steady load 2 operate automatically some early designs relied on the miners having to periodically strike the joint to initiate a degree of yield and 3 not snag on roadway lagging material during arch yielding in conclusion when comparing u k and german arch roadway support practices the following can be concluded performance at yield is clearly better for arches made of t h sections than for those made of h sections support dimensions are variable on setting for arches made of t h sections within limits they are not for arches made of h sections costs are higher for arches made of t h sections than for arches made of equivalent h sections mainly due to the clamps lateral strength is greater for t h sections and virtually nonexistent for h sections joints are strong points for arches made of t h sections due to overlap they are often the weakest points for arches made of h sections versatility is greater for h sections in that they can be readily adapted to u k needs to use square section or flat topped arches in in seam retreat roadways or r p districts the author is not aware of this need in germany or of any developments along this line using t h sections as a result of this mix of experiences and operational requirements the two industries have developed their roadway support technology as they have developed their mining systems along different lines face support longwall support offers many challenges to the engineer among its basic strata support capabilities a support unit must do the following be capable of being set at a variety of working heights impart an immediate positive load to the roof strata to prevent bed separation in the immediate roof strata and control the movement of bridging beds yield in a controlled manner within a chosen load range provide protective cover against falls of smaller pieces of rock into the working area support the coal face laterally prevent ingress of material or dust from the waste into the working area additionally the support unit must perform the following equally important operational tasks that are not necessarily linked to its ability to act as a roof support be capable of advancing the face conv

eyor and then moving the support structure forward be capable of retaining its alignment relative to its neighbors allow personnel access along the longwall have a low collapsed height for ease of transportation among longwall applications provide for both in support and remote control of support operations provide base positions for coal face lighting installations in the united kingdom investment in longwall plant and equipment has always been a major feature of the mining industry s capital expenditures second only to that for new mines or major reconstruction projects at existing mines by far the largest item in that expenditure is for powered roof supports because of their considerable cost their design features are of significant interest introduction of the lemniscate linkage to powered roofsupport designs is arguably one of the most influential technical achievements in longwall mining in modem times it unlocked the way toward the working of thicker seams in longwall systems in the united kingdom the major powered roof support manufacturers had developed chock support technology to its ultimate this was best demonstrated by the hugely successful six leg 240 t yield support with full shielding full width canopy and heavy duty forepole figure 8 7 13 although designed for extractions up to 2 4 m thick it was at its best in extractions of 2 m attempts had been made to develop chock support technology for extractions of 3 m but at this extraction height the stability of a 1 2 m wide support was questionable especially with regard to its lack of resistance to lateral loading due to its inability to operate successfully with coal face support sprags and to be in contact with its neighboring supports the lemniscate linkage changed the internal stability of the support structure it removed all lateral load from the support legs so that leg restraint need no longer be provided by leg towers supports could now be designed with their side shields and canopy side flaps in contact with those of their neighbors greatly reducing the ingress of material from the waste into the traveling track which had been a major problem with chock type supports supports could also be designed to advance while retaining contact with the roof a feature unheard of in conventional chock design the lemniscate linkage also addressed another problem that of spalling coal from the face a well known danger to workers that ultimately leads to exposure of unsupported roof at the face and increases the potential for broken ground or major falls the solution to this problem was the use of face sprags once again the lateral resistance offered by the lemniscate protects the legs from any resultant lateral loading lemniscate linkage supports became known as shield supports the first shield supports used in the united kingdom were four leg chock shields figure 8 7 14 operators traveled the longwall within the support structure behind the

front legs however although they offered positive support to the roof they did not have a particularly good working range inclined leg shields soon overtook them in all seam ranges they enabled the use of shield technology in thinner seams and allowed passage through the face in front of the forward legs providing better access to coal face equipment additionally the considerable range of working heights available with these types of support made them easy to transfer among different seams in the same mine and indeed different mines two leg shields were a natural extension of inclined leg designs figure 8 7 15 the rearward facing inclined legs were replaced by stabilizing rams between rear shield and support canopy interest in their use in the united kingdom developed as mining engineers became aware of the number of overseas applications individual leg ratings of 300 t fortwo leg shields led to a far superior density of support when compared with u k four leg designs however the chief argument that slowed their u k application was the perception that the forward edge of the base would dig into the soft floors encountered in many u k coal mines this valid concern was overcome by the introduction of the base lift which lifts the front end of the base during support advance how much support is required support ratings have increased over the last 30 years in terms of both the yield load by which a support is defined and the applied setting load the setting load has increased as a percentage of yield load in the united kingdom from about 30 to 75 however the question has never been satisfactorily answered as to how strong a support should be for a particular circumstance the support acquisition process in the united kingdom was based on the premise that the mine operator would select a support rating and the manufacturer would supply to that specification and demonstrate by means of rigorous laboratory testing that the support structure and hydraulic legs met that specification there is a history especially in europe that powered supports have tended to become overdesigned for their duty in terms of nominal load bearing capacity however in view of the substantial investment required to purchase supports the long arduous working life expected of them and the uncertainty always attendant upon soft rock mining it has been regarded as prudent to have a significant factor of safety when applying powered supports mechanical testing of support structures initially identified as a guide to design safety is now also used to measure design reliability such testing involves subjecting structures and hydraulic components to numbers of test cycles approaching those likely to be experienced during production life these requirements are customer led rather than imposed by regulatory bodies as the international market is populated by mine operating companies that are now well aware of what they require from a suppo

rt and how those features might be demonstrated by the potential supplier in 1999 mining regulations were introduced in the united kingdom to control the application of support in underground mines these constitute the first revision of the 1966 regulations that were drawn up when powered support designs were comparatively new before then a maximum interval was specified between the coal face and the nearest support member of any given support design this distance generally 2 m was taken as the measure of safe support and was the basic parameter that influenced face design especially the design of the power loader and the armored face conveyor and its furnishings even in an era of relatively low capacity face equipment this distance was barely adequate to house the face equipment as the power and size of all elements in a coal face package increased this propfree front distance plainly restricted potential production that could be won from the face some better measure of support had to be found early work by the mining research establishment ashwin et al 1970 demonstrated empirically the minimum mean load density mld required at the coal face to limit convergence during normal operations at mld 10 76 t m2 convergence differs very little from that at mld 10 76 t m2 at mld 10 76 t m2 convergence increases these optimum mld figures were later shown to vary with the seam section recommended load densities were then quoted as 11 8 t m2 for thin seams 22 6 t m2 for medium seams and 29 1 t m2 for thick seams up to a maximum of 2 4 m in 1982 the british coal corporation rewrote its internal owner rules governing the use of powered supports moving away from the single measurements mentioned previously their document set down for the first time instructions for determining the minimum setting and yield load densities required in the various zones of a coal face in the face zone minimum values were set to the following setting mld 7 5 h t m2 and yield mld 15 h t m2 where h is the seam section in meters thus a 1 m section requires a setting mld of 7 5 t m2 and a yield mld of 15 t m2 a 2 4 m section requires a setting mld of 7 5 t m2 and a yield mld of 37 5 t m2 the actual designed setting and yield densities for powered support have increased over the years reaching values well above these minimum requirements the focus of attention has since moved from yield density to setting density with setting densities now approaching within a few percentage points of yield densities the question of the validity of mld as the sole definition of a support s operational characteristic requires closer examination as it is an average figure its value must be qualified by reference to the distribution of the load on a support canopy and base however load distribution profiles are somewhat academic in concept because they relate to the ideal situation of a roof beam in full contact along its whole length with a

strata profile of matched shape a long way from normal operational conditions under which powered supports operate but it does demonstrate that for a support to operate effectively consideration must be given to not only mld but also canopy load distribution five and six leg chock supports because of their geometry exhibit a reasonably welldistributed load profile over canopy and base concern about maintaining an even load distribution in the base was one of the chief reasons that chock shield designs became so popular when the advantages of lemniscate supports were recognized by the industry the overall pattern of strata behavior around a longwall face has been widely researched since the early 1950s for virtually as long as full caving of waste and powered support have been in existence difficulties in making measurements in the strata above a working face have always dictated that information must be gathered from a number of sources namely observations of convergence and pressure profiles of the support legs in situ physical and mathematical modeling and pure hypothesis that the debate continued into the 1990s was symptomatic of the uncertainty that surrounded the subject when a stratified deposit is extracted by the longwall method the roof of the longwall is supported temporarily by powered supports until the system has advanced sufficiently that the supports can be removed and the roof allowed to freely collapse or cave behind them this volume of broken rock is known as waste the thickness of roof strata that collapses is governed by the type of strata that forms the roof soft friable rocks such as mudstones cave easily when unsupported breaking into relatively small fragments the volume of which is greater than the volume it originally occupied as intact rock this bulking process dictates the height to which caving can occur uncaved beds above the caving beds are known as the bridging beds these remain intact and subside over the caved waste in a manner controlled by the support offered by the unworked mineral ahead of the face the action of the powered supports and the compression of the caved material in the waste this process is known as bulking factor controlled caving gorrie and scott 1970 wilson 1975 b g d smart personal communication calculations show that bridging beds commence a distance of twice the height of the mineral extracted at the face above the face roof this gives an effective value to the weight of rock for which support must be designed convergence on a well supported longwall is caused by the downward movement of the bridging beds being transmitted through the immediate roof beds while they remain intact the support offered to the roof by the powered support maintains the integrity of the immediate roof beds but does little to prevent convergence of higher bridging beds when stronger rocks such as sandstone and limestone form the immediate roof of the mineral being worked on

a longwall they frequently do not cave immediately behind the powered supports but rather cantilever into the waste for some distance this puts additional load on the supports furthermore when this rock eventually caves it does not fragment as readily as does weaker material rather it forms large lumps in the waste with an effectively much lower bulking factor the distance up to the bridging beds thus becomes much larger beds can also cave to a well defined parting plane and the height of the parting plane above the mineral dictates the height of caving this phenomenon is called parting plane caving as a general rule the depth of a working from the surface does not directly affect the loading to be expected on the powered supports support loads experienced 200 and 1 200 m below the surface can be similar there is however one surprising exception to this rule perhaps the most difficult circumstances for longwall support come when working shallow 150 m depth seams with competent overlying strata caving in these circumstances often extends right through to the surface such that the powered supports are asked to support a column of strata from working level to surface in these circumstances supports designed to yield at 1 000 t are applied these mechanisms are derived from hypothesis rather than measurements for this reason engineers purchasing and installing powered supports for a new application apply a significant factor of safety to ensure that the supports do not at some time in their working lives become overloaded it has become usual in europe with its long history of using powered support for decisions on new support designs to be made empirically empirical selection of supports based on experience in comparable strata conditions rarely fails the chief criticism that it attracts is that supports can sometimes be overengineered this is not necessarily a bad practice as the extra cost of the overdesign which is difficult to precisely quantify serves as an insurance premium against the unknown in all but the thickest seams yield load densities in the united kingdom peak at 100 t m2 which normally relates to support yield loads of 650 750 t in the mid seam range the cost of achieving this density for the mine operator is high and a compromise must eventually be reached between the equally strong desires to increase loading and to restrict the cost of powered supports matching support size to an actual application still owes as much to experience as to science and economic reality the modern shield support has displayed considerable mechanical reliability and is well equipped to deal with unreliable or unknown geology however no amount of technology can maintain high production levels in the most faulted weak water bearing strata or where seam sections reduce the inflexibility of the longwall method will always make it vulnerable to uncertain geology standardization in the late 1980s the eur

opean union issued a directive covering the basic safety of machinery an amendment to that directive highlighted powered roof supports in mines as a type of machine that requires particularly careful treatment with respect to supply within the european community member states a working group comprised mostly of u k and german delegates representing mine operators the u k mines inspectorate and powered support suppliers put together a series of basic safety standards harmonizing european practice bs en 1804 series 2001 2006 however this standard does not cover the use of support underground safety with respect to powered supports is still judged largely by performance under test the major implications of this exercise are not to restrict the activities of u k and german support manufacturers but rather to control trading within the european union of potentially inferior products from suppliers outside the european union design methodologies in soft rock siddall and gale 1992 elegantly summarize a design philosophy appropriate to the identification of support systems to be used in soft rock environments in the following quotation the science of strata control would be best considered as the application of scientific principles to certain goals understanding the behavior of rock or strata under known conditions determining methods to control the strata under known conditions determining methods to predict the response of strata under conditions other than those for which current experience allows assessment determining methods of controlling the strata within limits dictated by mining operational viability in the conditions predicted outside those for which current experience allows assessment scientific principles cannot be applied unless the problems can be defined in a scientific format and the answers obtained can be seen to explain the actual site situation to achieve this process scientists and engineers need to obtain experience of the mine environment and understand the key factors influencing the behavior of the environment determine the factors which can be quantified and those which cannot be directly quantified and may need to be extrapolated or assessed qualitatively undertake measurement of the strata behavior and the influence of mining activity formulate conceptual models of the mechanics involved and the interrelationship of the variables present a number of key decisions are required when selecting a support type for a soft rock excavation the following key questions must be answered how much closure can be tolerated in the excavation this question presupposes that some amount of closure may on occasion be accommodated within the mining system this requirement suggests knowing not only the degree of final closure but also the patterns of closure related to both the passage of time and the continuation of working within the excavation to deal with closure without premature failure sup

port must incorporate an appropriate yield mechanism that is one that provides a level of resistance prior to yield that is largely maintained during the yielding process is support required for the roof sides and floor in softrock excavations support must be provided not just to the roof of the excavation but also to the sidewalls and on occasion to the floor what are the rock excavation formation methods in most soft rock excavations the time interval between rock excavation and setting of support is an important consideration early setting of permanent support or introduction of temporary support measures are essential to providing early resistance to movement of newly exposed rock faces this early application of support greatly influences the convergence process how significant is cost containment this is a key element in most soft rock mining systems as soft rock minerals tend to be relatively low in value and therefore production costs must be maintained within tight bounds the use of supports is a key element in mining costs and often compromises must be made in choosing between the best support type and an adequate but less expensive alternative in designing support systems for soft rock applications the following approaches are taken either singly or jointly experience support system design in soft rocks is based largely on previously proven successful practices from the same mine or neighboring mines with perceived similar conditions it is also often based on the personal experience of individual engineers this is a quite satisfactory approach and can provide safe and efficient support systems but it can also perpetuate overdesign of the support system that can prove ultimately expensive although it is quite correct to set additional support within a previously applied system for a new site only a brave or foolhardy engineer would decide to reduce support density in a mining system based purely on experience without recourse to measurement or rigorous design measurement a long tradition exists of measuring strata dilation around a mining excavation in its simplest form this is done by in tunnel roof to floor and side to side convergence tape measurements in general the degree of roof floor or side displacement can be measured only to the centimeter level and high levels of accuracy are unnecessary patterns of behavior can be identified and variations in the support applied can be evaluated by comparison of closure measurements with the introduction of rock bolted support in coal mine roadways in the united kingdom in the late 1980s to early 1990s a different approach to the design of roadway support was required rock bolts were a technology new to both mine management and miners and confidence had to be gained that these supports were adequate and safe this process coincided with a general renewal of u k health and safety legislation underway at the time a significant feature of the format o

f new mining regulations in the united kingdom was their lack of prescriptive detail in place of prescriptive detail were general principles and a stated requirement for mine operators to conduct risk assessments of all underground operations before commencement the approach adopted to safely introduce the new rockbolt support technology is to set down a process of geotechnical risk assessment largely based on site investigation leading to selection of a support system design key elements of this process are knowledge of the geology and stress field together with in situ experimentation to establish the bond strength between resin and strata at all major roof horizons within the bonded length of the bolt knowledge of the typical pattern of strata dilation in relation to its distance from the face once exposed is also needed the design of the support system is in practice empirically based and site specific taking into account certain general parameters such as minimum bolt length to be used 1 8 m minimum bolt density in the roof 1 bolt m2 hole diameter 7 mm larger than the bolt diameter and the need to place rock bolts as near to the face of the heading as possible and as soon as practicable after the roof is exposed support of the sides is now considered part of the overall design rock bolts after installation are subject to detailed and comprehensive monitoring by measurement of strata movement and rock bolt load extensometers are used to measure with a high degree of accuracy much smaller displacements experienced between different strata sections in the rock mass around the excavation operating rules also specify the presence of routine monitoring in every roadway in which rock bolting provided the principal support monitoring takes two forms dual height telltales and multi horizon extensometers dual height telltales are installed to at least twice the bolt length or 4 8 m whichever is greater at intervals of 20 m along the roadway additional telltales are required at every junction and other places where roof instability is known or suspected multihorizon extensometers reading ground movement to 7 m above the roof of the roadway are installed at maximum intervals of 200 m more importantly the mine manager is required to put into operation a scheme in which these routine monitoring stations are read and reported upon supervisors at the mine are given action levels of movement to measure against and actions to take if these levels are exceeded great britain health and safety commission 1996 physical modeling in the early years of strata control research a number of key mining research institutes studied the mechanisms of strata movement and stress redistribution around mining excavations especially longwall faces by physical modeling of mining situations using plaster models loaded in large frames this work was conducted in a number of coal mining countries and in its heyday was used to qualitativ

ely explain the mechanisms of displacement that take place away from the mining excavation these physical models were expensive in terms of resources and the results were often limited the models were heavily dependent on the size of the model domain the assumptions made and the scale effects for work of this type the results gave at best a broad qualitative view of large scale rockmass response to mining they were far less effective in providing practical data on which a support system can be designed numerical modeling before the computer era analytical solutions were developed that provided a conceptual understanding of rock mass behavior however in most cases they assumed rock to be a linear elastic material which is a large simplification especially in soft rock regimes for many situations they could solve only problems that had simple geometry and low dimensionality unfortunately rock masses are generally anisotropic inhomogeneous and inelastic due to the presence of geological discontinuities such as joints fissures and faults most softrock mining problems are nonlinear due to high levels of deformation and involve complex shapes and boundary conditions that often cannot be described by simple mathematical functions with the development of high speed personal computers in the last two decades powerful computer based numerical modeling techniques have become readily available for application to geotechnical materials numerical methods have become very popular in academia and more recently in industry for solving problems in rock mechanics the development of these numerical methods has provided the design engineer with a way to predict the stress displacement and failure mechanisms of zones adjacent to an excavation formed within rock masses that exhibit complex nonlinear constitutive behavior a number of modeling software products exist for solving geological and geotechnical problems both abaqus for finite element analysis and flac fast lagrangian analysis of continua software for finite difference analysis are popular for determining the stress conditions and behavior of rock masses surrounding an underground excavation flac has been used successfully for many underground problems in rock mechanics and is considered to be an industry standard in both its two dimensional figures 8 7 16 and 8 7 17 and three dimensional formats however no matter how good a model is a good prediction cannot be obtained unless the input data is of high quality unlike engineering materials rock mass generally contains a lot of irregular discontinuities in the form of joints bedding planes fissures and faults the engineering properties of in situ rock strata are determined by many factors such as intact rock properties nature and orientation of planes of weakness and presence of groundwater the stiffness and strength of intact rock can be determined in the laboratory providing some information about the uppe

r limit of the engineering properties of the rock mass however it is practically impossible to determine all of the properties of the many discontinuities by direct testing and measurement therefore the engineering properties of the rock mass may be estimated by means of a rock mass classification system to supplement direct measurement by testing fortunately several well recognized engineering rock mass classification systems have been widely used to estimate the strength and stiffness of rock mass including its discontinuities as a powerful aid in design rock mass classification has become an integral part of the empirical design methods popular in the industry and has been invaluable on many figure 8 7 16 typical flac output for stress distribution around a mine roadway figure 8 7 17 typical flac output for strata displacement around a mine roadway 610 sme mining engineering handbook underground construction and mining projects however rock mass classification systems alone are not the ultimate solution to design problems and they are now being applied in conjunction with numerical methods ground control is the art and science of maintaining safe and stable conditions at the surface of an excavation in a rock mass at or below the ground surface for the intended service life of the structure aggressive mining strategies design constraints on excavation geometry or unfavorable arrangements of rock stress structure or material properties often lead to the requirement for some form of rock mass improvement to maintain safe and stable conditions methods for achieving this improvement are given in numerous standard textbooks and conference proceedings e g hoek and bray 1974 hoek and brown 1980 stillborg 1986 stephansson 1983 the term support is widely used to describe procedures and materials that improve the stability and maintain the load bearing capability of rock near the boundaries of surface and underground excavations rock improvement includes all techniques that seek to increase the strength and decrease the deformability characteristics of the rock mass this includes the injection of chemical or cementitious grouts ground freezing presplitting and drainage within the general category of rock improvement rock reinforcement is a specific technique with the prime objective of improving the shear and tensile strength of the rock mass adjacent to surface and underground excavations the term support can then be reserved for use in those cases in which the rock mass is truly supported by structural elements that carry in whole or in part the weights of individual rock blocks isolated by individual discontinuities or zones of loosened rock brady and brown 2004 support and reinforcement are essential components in the design of all surface and underground excavations and are often combined to provide an overall or integrated system in mining they enable more economic extraction through steeper slopes in open p

its or increased recovery rates from underground operations in civil engineering they enable increased excavation rates in tunnels and safer permanent surface and underground facilities windsor and thompson 1993 in this chapter the dominant reinforcement and support elements will be discussed along with installation procedures and major quality assurance quality control qa qc issues initially however some basic terminology will be defined in an attempt to minimize confusion due to the divergent jargon that has evolved in the industry rock reinforcement techniques rock support versus reinforcement rock reinforcement is often mistakenly taken to include methods more properly termed support it is essential to the understanding of the underlying mechanics to differentiate between the concepts of reinforcement and support the primary objective of rock mass reinforcement is to mobilize and conserve the inherent strength of the rock mass so that it becomes selfsupporting ground support on the other hand should be restricted to structural elements that directly carry the weight or part thereof of discrete rock blocks support and reinforcement should also be classified as active or passive active support imposes a predetermined load to the rock surface during installation and can take the form of tensioned rock bolts or cables hydraulic props expandable segmented concrete linings or powered support for longwall faces active support is usually required to support the gravity loads imposed by individual rock blocks or loosened zones a substantial body of work has been undertaken in the australian coal industry to demonstrate that active support via bolt and cable pre tension is a critical aspect of strata reinforcement where the objective is to control in situ ground stresses rather than the weight of potentially unstable strata passive support or reinforcement is not installed with an applied loading but develops its loads as the rock mass deforms in fact many reinforcement schemes only operate when the rock mass reaches yield passive support may be provided by steel arches timbered sets or composite packs or by untensioned grouted rock bolts reinforcing bars or cables these three often described as dowels strata control describes the support and reinforcement techniques used in coal mining although the term is suitable because it evokes the concept of the control or limitation of strata displacements rather than support the major function of some strata control measures is support brady and brown 2004 load transfer concepts for reinforcement fundamental to the understanding of rock reinforcement behavior and the action of different elements and their effect on excavation stability is the load transfer concept this is shown schematically in figure 8 8 1 and is composed of three basic mechanisms 1 rock movement that requires load transfer from the unstable rock to the reinforcing element 2 transfer of load via t

he reinforcing element from the unstable surface region to the stable interior region 3 transfer of the reinforcing element load to the stable rock mass reinforcing element length can be considered in three general ranges associated with perceived unstable surface regions for different surface and underground excavations near surface 1 5 3 0 m long medium depth 3 15 m long and deep seated 15 m long windsor and thompson 1993 load transfer can occur by numerous methods and many different reinforcing devices have been developed in general these devices are solid bar hollow tube or wire rope elements that are installed within boreholes drilled into the rock a terminology has been proposed windsor and thompson 1993 that places load transfer mechanisms for these devices within one of three categories as defined in figure 8 8 2 these categories are 1 continuously mechanically coupled cmc 2 continuous frictionally coupled cfc and 3 discrete mechanically and frictionally coupled dmfc continuous mechanically coupled elements these elements rely on a securing agent that fills the annulus between the element and the borehole wall generally known as grout the agent is placed as a fluid and requires a finite setting time before the element is able to provide support the setting time can vary from a few seconds to several hours depending on the type of grout used which may well affect selection of the device the major function of the grout is to provide a mechanism for load transfer between the rock and the reinforcing element cmc in figure 8 8 2 reinforcing elements used with grouts are often manufactured with variable cross sectional shapes this variation causes a mechanical interlock between the element and the grout and strongly impacts the shear resistance or bond strength and the loadtransfer stiffness that can develop at the element grout interface continuous frictionally coupled elements cfc elements behave similarly to cmc elements except the cfc element is placed in direct contact with the rock cfc in figure 8 8 2 load transfer results from friction between the element and the borehole and is limited by the radial prestress set up at installation any geometrical key that is present occurs by chance due to irregularities in the borehole wall devices in this class consist either of expansion of an undersized section into a larger borehole or contraction of an oversized section into a smaller borehole see frictional support section borehole size is critical to support performance in this class discrete mechanically and frictionally coupled elements dmfc elements tend to be simple bars with either a deformed end region or provision at one end for attaching an expanding anchor and provide either mechanical or frictional load transfer over this relatively short portion of the entire element length dmfc in figure 8 8 2 this interval is commonly known as the anchorage length and is usually limite

d to less than 500 mm for grouted anchorages and 200 mm for expansion anchorages the anchorage must have sufficient capacity to mobilize the full material strength of the element expansion shell bolts are the most common dmfc element see the ungrouted mechanically anchored rock bolts section with these elements the anchor is expanded outward against the borehole wall because of tension in the element and anchorage strength may be limited by the strength of the host rock grouted anchorages have a lower unit load transfer and offer advantages in weaker and softer rocks in general only resinous grout can provide the high strength requirements for shorter anchorages pre and postexcavation reinforcement and support pre tensioning in certain applications it may be advantageous to install support elements prior to excavation i e pre reinforcement while in other cases it may be better to install the support elements at some discrete time following excavation in the former case support elements may better restrain rock mass movements but may be required to sustain higher loads in the latter case ultimate support element loads can be controlled by allowing some initial rock movement this is common in tunneling using concrete linings or shotcrete in weak strata where one objective is to ensure the long term integrity of the permanent support e g the new austrian tunnelling method rock support interaction theory provides a framework for understanding the impact of pre versus postexcavation practice and is covered in numerous standard rock mechanics textbooks e g brady and brown 2004 hoek et al 1995 the limitations of rock support interaction theory however must be clearly understood particularly that subsequent extraction processes change the ground stresses which therefore may require higher levels of support than development conditions in certain applications it is also advantageous to provide the reinforcing element with an initial pre tension posttensioning is the tensioning or retensioning of elements subsequent to installation it is important to understand that additional tension is likely to develop in the element because of rock mass deformations in response to subsequent excavation stress change creep and so forth and to allow for this in order to avoid subsequent overstress and rupture of the support elements windsor and thompson 1993 encapsulation techniques encapsulation materials can generally be classified as cementitious usually normal portland cement or resinous with the latter generally being stronger cement grouts are suited to longer reinforcing elements because of availability ease of mixing and placement resins are restricted to shorter holes primarily because of difficulties in mixing resins at lengths exceeding 3 4 m a benefit of encapsulation is increased resistance to corrosion although in highly corrosive environments additional protection beyond simple encapsulation is req

uired resin grouted reinforcement has found wide application in discrete load transfer elements set times can be reduced to less than 10 seconds which is particularly important for cyclic and fast advance mining methods where support installation is one of the controls on advance rate it must be understood however that many other factors e g machine availability operator training and cutting rates also strongly influence advance rates in addition the anchors can supply design capacity within a few minutes of installation which improves safety during advance high early strength grouts are also available in cartridges windsor and thompson 1993 auxiliary fittings various forms of auxiliary fittings have evolved for use with different reinforcing elements that cater to different geological conditions the fittings are either attached to the reinforcing element at the rock surface or are used to modify the internal response of cmc elements special equipment and procedures are required for tensioned applications external fittings e g plates straps mesh can be attached to the reinforcing element to provide varying degrees of surface restraint to the rock mass external fittings are mandatory for discrete coupled devices while internal fittings are used primarily with cmc devices and may comprise intermittent anchors and or decoupling sleeves these combinations of devices can be used to engineer the stiffness of the reinforcing elements to match anticipated rock mass deformations based on predicted failure mechanisms in most applications the number of reinforcing elements in relation to the exposed excavation area is quite low in closely jointed fissile or weathered rock masses where unraveling may be the dominant failure mode stability can be improved by installing structural elements that span the reinforcing elements such support forms a surface retention system that restrains the movement of small blocks and provides a reactive stress to the free span area this stress is usually low for mesh applications but will be higher for shotcrete the concept of surface retention support on a larger scale using reinforcement as anchors is important and holds the key to supporting many of the larger more complex failure modes particularly for surface excavations windsor and thompson 1993 static versus dynamic support the majority of ground support applications are designed for static rock mass conditions which may include both simple gravity driven e g wedge failure and unraveling and mineinduced stress driven failure in the latter case a stable failure process occurs where most of the stored strain energy in the rock mass is consumed during the fracture and deformation process ground support must be designed with sufficient capacity for both anticipated loads and deformations however these will accumulate over time and energy absorption will not be a concern dynamic support elements are generally restricted

to ground subject to rock burst loading first are some basic definitions kaiser et al 1996 a seismic event is a transient earth motion caused by a sudden release of potential or stored strain energy in the rock as a result seismic energy is radiated in the form of strain waves the magnitude of a seismic event is usually determined from the peak amplitude of the strain wave using a logarithmic scale e g nuttli or richter scale a rock burst is a seismic event that causes injury to persons or damage to underground workings the general and essential feature of rock bursts is their sudden violent nature rock bursts are generally divided into three classifications 1 strain bursts these are caused by high stress concentrations at the edge of mine openings that exceed the strength of the rock events can range from small slivers of rock being ejected from the walls to the collapse of a complete wall as it tries to achieve a more stable shape 2 pillar bursts severe rock bursts involving thousands of tons have been caused by the complete collapse of support pillars these tend to occur in extensively mined out areas and the resulting damage can be severe 3 fault slip bursts recognized in the 1980s this type of rock burst occurs when slippage suddenly occurs along a geological weakness plane this is the same mechanism as for an earthquake rock burst magnitude and hence energy and damage potential increases from classifications 1 to 3 the damage may occur as rock bulking due to fracturing rock ejection by seismic energy or rockfalls due to seismic shaking figure 8 8 3 each mechanism may result in different levels of damage to an excavation and its support system the damage severity depends on many factors including failure potential near the opening i e the level of existing wall stress compared to the rock mass strength which in turn depends on rock wall quality support effectiveness local rock stiffness magnitude of seismically induced stress rock accelerations or velocities opening geometry size and orientation and geological structure three levels of rock burst damage severity are defined as shown in figure 8 8 4 damage level determination is usually based on observations of previous damage where such observations are available plus analytical methods the three severity levels are discussed later and are summarized in table 8 8 1 only a limited number of ground support elements are suitable for dynamic loading conditions see spiling section and under the most severe conditions even these are restricted by maximum practical support limitations rock reinforcement support elements rock reinforcement elements can be obtained in various lengths the most common are near surface support elements generally 3 m long that provide support to the inner shell of the opening and are commonly referred to as primary support deeper ground reinforcement can only be achieved either b

y using cable bolts or by connecting shorter stiff bolt elements together as discussed later in this section such deep reinforcement is commonly referred to as secondary support near surface static reinforcement 3 m depth ungrouted mechanically anchored rock bolts mechanically anchored reinforcement elements commonly referred to as point anchor bolts or expansion shell bolts fall into the dmfc category and represent one of the earliest developments in rock mass reinforcement these elements apply a positive force to the rock without requiring movement of the rock mass they are composed of three elements 1 an expansion shell anchor 2 a faceplate assembly and 3 a connecting steel bar expansion shell anchors consist of a tapered cone with internal thread plus a pair of wedges held with bail wire the cone is screwed onto the threaded end of the bolt and the assembly is inserted into the borehole the hole must be at least 100 mm longer than the bolt or the bail wire will be dislodged and the anchor will not set once in place the anchor is seated with a sharp pull on the bar tightening the bolt forces the cone into the wedge increasing the anchor force expansion shell anchors can work well in hard rock but are not good in closely jointed or soft rocks because of the potential for deformation and failure of the rock in contact with the grips at the bolt head either a fixed head forged head or threaded end and nut system can be used a faceplate is always required to distribute load from the bolt to the rock face a tapered washer or conical seat is required to compensate for the rock surface not being perpendicular to the bolt i e tightening the nut must place the bolt in direct axial tension a wide variety of faceplates and domed or tapered washers are available details of these assemblies are available from the various bolt manufacturers the bar element in a mechanically anchored rock bolt is a smooth circular element steel grade and bar diameter can be varied to adjust reinforcement element capacity and stiffness grouted mechanically anchored rock bolts mechanically anchored bolts can be changed from dmfc to a cmc category by filling the annulus between the steel element and the borehole wall with grout cement grout is used for this purpose and can be emplaced by the use of grout tubes inserted with the bolt figure 8 8 5 or through the use of a hollow stem bolt grout tubes are not commonly used with mechanically anchored bolts since space constraints restrict the grout tube diameter and therefore the water cement w c ratio of the grout that can be used a w c 0 4 grout should be used with a hollow stem bolt grout is pumped through the hollow core of the bolt and is forced back through the annulus grouting mechanically anchored bolts provides corrosion protection stops anchor slip and ensures that the full bolt capacity can be achieved hollow core bolts are available from a number of manufacturers

quality assurance and control qa qc is critical for all rock reinforcement elements tensioning of rock bolts is critical to ensure all components are in contact and that a positive force is applied to the bolt where bolts are required to carry significant load they are generally tensioned to about 70 of yield this provides a known load with reserve for additional load induced by rock movements corrosion is one of the primary causes of bolt failure this can be counteracted by grouting the bolts the use of hollow stem bolts e g stelpipe bolt makes this function simple to accomplish it is nevertheless a two stage process and therefore impacts cost the primary purpose in grouting the bolts is corrosion protection and to lock the anchor in place to stop loosening from blasting grout strength is important and higher w c grout can be used i e in the region of 0 4 w c ungrouted bolts must be retensioned after each blast for several rounds behind the face since blast vibration will cause anchor slip and loss of prestress a torque wrench can be used to test bolt tension beveled or conical seat washers must be used to ensure that tensioning of the bolt results in increased axial tension in the bar and subsequent seating of the anchor otherwise for applications not perfectly perpendicular to the rock face high torque on the nut will result from friction between the nut and the plate rather than tension in the bar resin anchored rebar bolts pre tensioned bolts ungrouted mechanically anchored bolts can work loose when subjected to blast vibrations or when anchored in weak rock additionally they are subject to corrosion and provide only relatively soft reinforcement resin anchored rebar overcomes these difficulties pre tensioned rock bolt applications use two component cartridges containing a resin and catalyst in separate compartments figure 8 8 6 the resin cartridges are inserted in the drill hole ahead of the bolt element which is then spun into them mixing the catalyst and the resin setting of the resin occurs in a few minutes depending on mix specifications and a strong anchor is created the system works for most rock types including weak rocks fully resin bonded bolts should be used for corrosion protection and long term support pre tensioning of the reinforcing element is achieved by using a combination of slow and fast setting resins in a pre tensioning application a fast setting resin 10 second set time is inserted first followed by slow setting resin e g 10 minutes for the remainder of the hole the bolt can be tensioned 2 to 3 minutes after installation the bar element is a deformed reinforcing bar with the toe end cut at about 45 degrees so as to adequately puncture the resin cartridges the high resin unit cost is offset by the installation speed of the tensioned fully grouted bolt dowels dowels are rock reinforcement elements installed passively in a borehole i e no prestress such el

ements require movement of the rock mass in order to develop a restraining force dowels work well when applied close to the advancing face and or where later mine induced stress change will cause movement of the rock mass the most common form is a resin grouted dowel in north america this consists of a resin grouted rebar dowel where single set time resin is used for full dowel encapsulation resin grouted dowels and rock bolts are in the cmc category of support resingrouted rebar forms a very stiff support the capacity of which is dictated by the type of steel and the rebar diameter quality assurance and control quality control issues with resin rebar support primarily revolve around problems with the resin resin qa qc problems include the following shelf life is limited may be as short as 6 months resins must be stored under conditions conforming to manufacturer s recommendations including storage temperature and conditions freezing or excessive temperature may impact resin characteristics effective mixing of the resin can be a problem cutting the bolt rod at an angle to form a sharp point will help properly puncture the resin cartridges the allowable length of time or number of rotations for spinning the resin is limited additional spinning destroys the resin structure and weakens the overall installation follow the manufacturer instructions in this area hole diameter is critical because it exerts a strong control on bond strength and load transfer stiffness in some weak argillaceous rocks the drill hole surface can become clay coated and the resin will spin in the hole causing incomplete mixing the resin initially has a low viscosity and in highly fractured rock the resin may seep into fractures leaving voids in resin column the reaction of resins with aggressive groundwater is not certain frictional support split set bolts split set stabilizers consist of a slotted hollow high strength steel tube and faceplate figure 8 8 7 the bolts are installed by pushing them into a slightly undersized hole the radial spring force generated by compression of a c shaped tube provides frictional anchorage along the complete length of tube the toe end of the tube is beveled inward to facilitate installation and the system is quick and easy to install it can be particularly useful in mild rock burst conditions where the bolt will slip and not break and when used with mesh will retain broken rock generated by a mild burst these bolts can also be used to pin up mesh in highly fractured and yielded ground prior to the application of shotcrete provided the load demand does not exceed capacity this system can work well for short term support however susceptibility to corrosion rules out this system for most longterm support requirements the bolts are installed as dowels and fall in the cmc category table 8 8 2 provides typical split set bolt specifications these may vary for different manufacturers

split set bolt pull out capacity is rated in metric tons per meter t m and is derived from in situ pull tests nominal pull out strengths vary from 1 to 2 t m swellex bolts the original swellex dowel came in varying lengths and consisted of a 42 mm diameter tube that is folded into a 25 28 mm diameter during manufacture when the dowel is inserted into a 32 39 mm drill hole no push force is required the dowel is activated by injection of high pressure water about 30 mpa which inflates the folded tube into intimate contact with the walls of the drill hole in 1993 the original swellex dowel was replaced with the exl swellex manufactured from high strength but ductile steel present products consist of the swellex premium pm line table 8 8 3 and the swellex manganese mn line table 8 8 4 of bolts the bolts are available in various lengths up to 7 m and in coated versions for corrosion protection one of the most important features of the swellex system is the high bolt ductility the pm line of bolts can withstand 10 strain while the mn line can withstand 20 strain this high ductility without loss of capacity makes the swellex bolt useful in rock burst and high closure applications an additional advantage of this system is the speed of installation compared to conventional rock bolts and cement grouted dowels quality assurance and control split set bolts corrosion is a potential problem with all split set bolts because of the thin steel section of the tube that is in direct contact with the rock although galvanizing the tube helps delay corrosion much of the galvanizing may be scraped off during installation and the impact of this on bolt life is uncertain these bolts are generally not suitable for long term support and should not be used in areas with aggressive groundwater split set bolts have limited capacity and type ss 39 will only mobilize about 5 t on pullout in many cases actual field loading conditions will reduce active capacity to no more than 50 of the pullout load the bolt capacity is sensitive to hole gauge and the bolts have extremely low shear resistance due to the limited steel cross sectional area swellex bolts corrosion of swellex dowels is of concern because the outer surface is in direct contact with rock and the steel section of the tube is thin coated swellex bolts are now available that help to reduce corrosion susceptibility the swellex coatings should be less impacted by installation procedures than is the case with galvanized split set bolts the long term support characteristics of swellex bolts particularly in aggressive groundwater conditions are nevertheless unknown swellex dowels can be tested for corrosion by repressurizing the bolts if the bolt holds the pressure it has not been perforated by corrosion these bolts are also sensitive to hole gauge and are weak in shear due to the limited steel cross sectional area spiling spiling is a method of advancing a tunnel in coll

apsing soil or very loose rock e g poor ground fault gouge lowangle holes are drilled ahead of the face using a sacrificial drill bit and hollow rods that may be coupled to achieve the desired length grout is then injected through the hollow rods to consolidate the ground between the closely spaced anchors forming a reinforced umbrella or diaphragm and the next tunnel round is then advanced below this support following the advance additional ground support is installed and another set of anchors is drilled and grouted and the tunnel is again advanced the self drilling anchors are critical to this process numerous bit sizes and styles are available for varying conditions grout may be injected through perforations in the bit or the hollow rods may also be perforated there are numerous manufacturers of such systems near surface static design guidelines ground support design is a complex topic and a specialist area the details of which are well beyond the scope of this chapter at the simplest level for static design there are empirical design techniques that are relatively easy to apply providing one fully understands the limitations of the supporting database and associated design methodology and examples of these are provided in this section the most basic design uses rules of thumb such as those provided by the u s army corps of engineers table 8 8 5 empirical ground support design charts have also been developed around rock mass classification charts e g the rock mass rating system and tunnelling rock mass quality index q system see chapter 8 4 figure 8 8 8 shows the chart for estimating support categories based on tunneling quality index q palmstrom and broch 2006 more sophisticated support designs incorporate various numerical modeling techniques e g unwedge phase2 flac3d discussion of these techniques is beyond the scope of this chapter near surface dynamic reinforcement elements near surface reinforcement elements suitable for use under dynamic loading include cone bolts swellex bolts and split set bolts cone bolts were originally developed in south africa and were later modified by noranda technology for use in north america as a resin bolt the mcb33 modified cone bolt for 33 mm holes cone bolt is a smooth bar threaded at one end with forged cone and mixing blade at the other end and is manufactured and distributed by mansour mining in sudbury canada it is installed in a specially formulated coneloc or fasloc resin and is mixed by the resin mixing blade a plastic sleeve is installed over the shaft of the bolt to debond the resin from the bar thus allowing the cone to yield or plough through the resin the cone bolt load displacement characteristic curve is shown in figure 8 8 9 cone bolts show high ductility and energy absorption characteristics cone bolts should provide high shear resistance swellex bolts have high ductility 10 20 strain assuming they are anchored beyond