Rock Slope Engineering Civil and mining 4th edition phần 8 ppsx

298 Stabilization of rock slopes depth, and that there will be no loss of load with time. A suitable testing procedure has been drawn up by the Post Tensioning Institute (1996) that comprises the following four types of tests: (a) Performance test; (b) Proof test; (c) Creep test; and (d) Lift-off test. The performance and proof tests consist of a cyc- lic testing sequence, in which the deflection of the head of the anchor is measured as the anchor is tensioned (Figure 12.11). The design load should not exceed 60% of the ultimate strength of the steel, and the maximum test load is usually 133% of the design load, which should not exceed 80% of the ultimate strength of the steel. As a guideline, performance tests are usually car- ried out on the first two to three anchors and on 2% of the remaining anchors, while proof tests are carried out on the remainder of the anchors. The testing sequences are as follows, where AL is an alignment load to take slack out of the anchor assembly and P is the design load (Figure 12.12(a)): Performance test: AL, 0.25P AL, 0.25P, 0.5P AL, 0.25P, 0.5P, 0.75P AL, 0.25P, 0.5P, 0.75P, 1.0P AL, 0.25P, 0.5P, 0.75P, 1.0P, 1.2P, AL, 0.25P, 0.5P, 0.75P, 1.0P, 1.2P, 1.33P—hold for creep test ∗ AL, P—lock off anchor, carry out lift-off test. Proof test: AL, 0.25P, 0.5P, 0.75P, 1.0P, 1.2P, 1.33P—hold for creep test ∗ P—lock off anchor, carry out lift-off test. ∗ Creep test—elongation measurements are made at 1, 2, 3, 4, 5, 6 and 10 minutes. If the total creep exceeds 1 mm between 1 and 10 minutes, the load is maintained for an addi- tional 50 minutes with elongation measurements made at 20, 30, 40, 50 and 60 minutes. The usual method of tensioning rock bolts is to use a hollow-core hydraulic jack that allows the applied load to be precisely measured, as well as cycling the load and holding it constant for the creep test. It is important that the hydraulic jack be calibrated before each project to ensure that the indicated load is accurate. The deflection of Figure 12.11 Test set-up for a tensioned multi-strand cable anchor comprising hydraulic jack with pressure gauge to measure load, and dial gauge on independent mount to measure anchor elongation. (Photograph by W. Capaul.) Stabilization of rock slopes 299 1 0 0.25P 0.50P 0.75P 1.00P 1.20P 1.33P AL 2 3 4 5 6 10 min. Movement  t  e  r  t  e  t Load Residual movement Elastic movement Load Acceptance criteria 6 6 5 5 4 4 3 3 2 2 1 1 0 AL 0.50P 0.75P 0.25P 1.00P 1.20P 1.33P Line B: unbonded length + 50% bond length Line A: 80% free length (a) (b) Figure 12.12 Results of performance test for tensioned anchor: (a) cyclic load/movement measurements; (b) load/elastic movement plot (PTI, 1996). the anchor head is usually measured with a dial gauge, to an accuracy of about 0.05 mm, with the dial gauge mounted on a stable reference point that is independent of movement of the anchor. Figure 12.11 shows a typical test arrangement for tensioning a cable anchor comprising a hydraulic jack, and the dial gauge set up on tripod. The purpose of the performance and creep tests is to ensure that the anchor can sustain a con- stant load greater than the design load, and that the load in the anchor is transmitted into the rock at the location of the potential slide surface. The creep test is carried out by holding the maximum test load constant for a period up to 10 minutes, and checks that there is no significant loss of load with time. The creep test also removes some of the initial creep in the anchor. The lift-off test checks that the tension applied during the test- ing sequence has been permanently transferred to the anchor. The Post Tensioning Institute (PTI) 300 Stabilization of rock slopes 0.36 mm 1–10 minutes 0 36.0 36.1 36.2 36.3 36.4 36.5 36.6 36.7 36.8 36.9 37.0 37.1 1234 Log time (minutes) 56 10 Acceptance criterion (1 mm) Movement (mm) Figure 12.13 Results of creep test showing measured elongation over 10 minutes test period compared with acceptance criteria of 1 mm elongation. provides acceptance criteria for each of the four tests, and it is necessary that each anchor meets all the acceptance criteria. The results of a performance test shown in Figure 12.12(a) are used to calculate the elastic elongation δ e of the head of the anchor. The total elongation of the anchor during each load- ing cycle comprises elastic elongation of the steel and residual δ r (or permanent) elongation due to minor cracking of the grout and slippage in the bond zone. Figure 12.12(a) shows how the elastic and residual deformations are calculated for each load cycle. Values for δ e and δ r at each test load, together with the PTI load–elongation acceptance criteria, are then plotted on a separate graph (Figure 12.12(b)). For both performance and proof tests, the four acceptance criteria for tensioned anchors are as follows: First, the total elastic elongation is greater than 80% of the theoretical elongation of the unbonded length—this ensures that the load applied at the head is being transmitted to the bond length. Second, the total elastic elongation is less than the theoretical elongation of the unbon- ded length plus 50% of the bond length—this ensures that load in the bond length is con- centrated in the upper part of the bond and there is no significant shedding of load to the distal end. Third, for the creep test, the total elonga- tion of the anchor head during the period of 1–10 minutes is not greater than 1 mm (Figure 12.13), or if this is not met, is less than 2 mm during the period of 6–60 minutes. If necessary, the duration of the creep test can be extended until the movement is less than 2 mm for one logarithmic cycle of time. Fourth, the lift-off load is within 5% of the designed lock-off load—this checks that there has been no loss of load during the operation of setting the nut or wedges, and releasing the pressure on the tensioning jack. The working shear strength at the steel–grout interface of a grouted deformed bar is usually Stabilization of rock slopes 301 greater than the working strength at the rock– grout interface. For this reason, the required anchor length is typically determined from the stress level developed at the rock–grout interface. 12.4.3 Reaction wall Figure 12.4, item 3 shows an example where there is potential for a sliding type failure in closely fractured rock. If tensioned rock bolts are used to support this portion of the slope, the frac- tured rock may degrade and ravel from under the reaction plates of the anchors, and eventu- ally the tension in the bolts will be lost. In these circumstances, a reinforced concrete wall can be constructed to cover the area of fractured rock, and then the holes for the rock anchors can be drilled through sleeves in the wall. Finally, the anchors are installed and tensioned against the face of the wall. The wall acts as both a pro- tection against raveling of the rock, and a large reaction plate for the rock anchors. Where neces- sary, reinforced shotcrete can be substituted for concrete. Since the purpose of the wall is to distribute the anchor loads into rock, the reinforcing for the wall should be designed such that there is no cracking of the concrete under the concentrated loads of the anchor heads. It is also important that there are drain holes through the concrete to prevent build-up of water pressure behind the wall. 12.4.4 Shotcrete Shotcrete is a pneumatically applied, fine- aggregate mortar that is usually placed in a 50–100 mm layer, and is often reinforced for improved tensile and shear strength (American Concrete Institute, 1995). Zones and beds of closely fractured or degradable rock may be pro- tected by applying a layer of shotcrete to the rock face (Figure 12.4, item 4). The shotcrete will control both the fall of small blocks of rock, and progressive raveling that could eventually produce unstable overhangs. However, shotcrete provides little support against sliding for the overall slope; its primary function is surface pro- tection. Another component of a shotcrete install- ation is the provision of drain holes to prevent build-up of water pressures behind the face. Reinforcement. For permanent applications, shotcrete should be reinforced to reduce the risk of cracking and spalling. The two common meth- ods of reinforcing are welded-wire mesh, or steel or polypropylene fibers. Welded-wire mesh is fab- ricated from light gauge (∼3.5 mm diameter) wire on 100 mm centers, and is attached to the rock face on about 1–2 m centers with steel pins, com- plete with washers and nuts, grouted into the rock face. The mesh must be close to the rock surface, and fully encased in shotcrete, taking care that there are no voids behind the mesh. On irregular surfaces it can be difficult to attach the mesh closely to the rock. In these circumstances, the mesh can be installed between two layers of shotcrete, with the first layer creating a smoother surface to which the mesh can be more readily attached. An alternative to mesh reinforcement is to use steel or polypropylene fibers that are a compon- ent of the shotcrete mix and form a reinforce- ment mat throughout the shotcrete layer (Morgan et al., 1989, 1999). The steel fibers are manu- factured from high strength carbon steel with a length of 30–38 mm and diameter of 0.5 mm. To resist pullout, the fibers have deformed ends or are crimped. The proportion of steel fibers in the shotcrete mix is about 60 kg/m 3 , while compar- able strengths are obtained for mixes containing 6 kg of polypropylene fibers per cubic meter of shotcrete. The principal function of fibers is to significantly increase the shear, tensile and post- crack strengths of the shotcrete compared to non-reinforced shotcrete; shotcrete will tend to be loaded in shear and tension when blocks of fractured rock loosen behind the face. The disadvantages of steel fibers are their tend- ency to rust at cracks in the shotcrete, and the hazard of the “pin cushion” effect where per- sons come in contact with the face; polypropylene fibers overcome both these disadvantages. 302 Stabilization of rock slopes Mix design. Shotcrete mixes comprise cement and aggregate (10–2.5 mm aggregate and sand), together with admixtures (superplasticizers) to provide high early strengths. The properties of shotcrete are enhanced by the use of micro-silica that is added to the mix as a partial replacement for cement (USBM, 1984). Silica fume is an ultra fine powder with a particle size approximately equal to that of smoke. When added to shotcrete, silica fume reduces rebound, allows thicknesses of up to 500 mm to be applied in a single pass, and covers surfaces on which there is running water. There is also an increase in the long-term strength in most cases. Shotcrete can be applied as either a wet-mix or a dry-mix. For wet-mix shotcrete the compon- ents, including water, are mixed at a ready-mix concrete plant and the shotcrete is delivered to the site by ready-mix truck. This approach is suitable for sites with good road access and the need for large quantities. For dry-mix shotcrete the dry components are mixed at the plant and then placed in 1 m 3 bags that have a valve in the bottom (Figure 12.14). At the site, the bags are discharged into the hopper on the pump and a pre-moisturizer adds 4% water to the mix. The mix is then pumped to the face where additional water is added through a ring valve at the nozzle. The advantages of the dry-mix process are its use in locations with difficult access, and where small quantities are being applied at a time. It is also useful to be able to adjust the quantity of water in areas where there is varying amounts of seepage on the face. Typical mixes for dry-mix and wet-mix silica fume, steel fiber reinforced shotcrete are shown in Table 12.9 (Morgan et al., 1989). Shotcrete strength. The strength of shotcrete is defined by three parameters that correspond to the types of loading conditions to which shotcrete may be subjected when applied to a slope. Typical values for these parameters are as follows: (a) Compressive strength of 20 MPa at 3 days and 30 MPa at 7 days; (b) First crack flexural strength of 4.5 MPa at 7 days; and (c) Toughness indices of I 5 = 4 and I 10 = 6. Figure 12.14 Dry-mix shotcrete process using bagged mix feeding a pump and pre-moisturizer. Stabilization of rock slopes 303 Table 12.9 Typical shotcrete mixes Material Dry-mix Shotcrete Wet-mix Shotcrete (kg/m 3 ) (% dry materials) (kg/m 3 ) (% dry materials) Cement, Type I 400 18.3 420 18.3 Silica fume 50 2.3 40 1.7 10 mm coarse aggregate 500 22.9 480 20.9 Sand 1170 53.7 1120 48.7 Steel fibers 60 2.8 60 2.6 Water reduced — — 2l 0.09 Superplasticizer — — 6l 0.04 Air entraining admixture — — if required if required Water 170 a — 180 7.8 Total wet mass 2350 100 2300 100 Note a Total water from pre-moisturizer and added at nozzle (based on saturated surface dry aggregate concept). Toughness indices Load A C E FD Deformation B0  3 5.5 1234 l 5= A OAB A OACD l 10 = A OAB A OAEF Figure 12.15 Load–deformation characteristics of steel-fiber reinforced shotcrete. 1, without fibers; 2, 1% vol. fibers; 3, 2% vol. fibers. 4, 3% vol. fibers (American Concrete Institute, 1995). The flexural strength and toughness indices are determined by cutting a beam with dimensions of 100 mm square in section and 350 mm long from a panel shot in the field, and testing the beam in bending. The test measures the deforma- tion beyond the peak strength, and the method of calculating the I 5 and I 10 toughness indices from these measurements is shown in Figure 12.15. Surface preparation. The effectiveness of shotcrete is influenced by the condition of the rock surface to which it is applied—the sur- face should be free of loose and broken rock, soil, vegetation and ice. The surface should also be damp to improve the adhesion between the rock and the shotcrete, and the air temperature should be above 5 ◦ C for the first seven days when the shotcrete is setting. Drain holes should be drilled through the shotcrete to prevent build- up of water pressure behind the face; the drain holes are usually about 0.5 m deep, and located on 1–2 m centers. In massive rock the drain holes should be drilled before the shotcrete is applied, and located to intersect discontinuities that carry water. The holes are temporarily plugged with wooden pegs or rags while applying the shotcrete. 304 Stabilization of rock slopes Aesthetics. A requirement on some civil pro- jects is that shotcreted faces should have a natural appearance. That is, the shotcrete should be colored to match the natural rock color, and the face sculpted to show a pattern of “discon- tinuities.” This work is obviously costly, but the final appearance can be a very realistic replica of a rock face. 12.4.5 Buttresses Where a rock fall or weathering has formed a cavity in the slope face, it may be necessary to con- struct a concrete buttress in the cavity to prevent further falls (Figure 12.4, item 6). The buttress fulfills two functions: first, to retain and protect areas of weak rock, and second, to support the overhang. Buttresses should be designed so that the direction of thrust from the rock supports the buttress in compression. In this way, bending moments and overturning forces are eliminated and there is no need for heavy reinforcement of the concrete, or tiebacks anchored in the rock. If the buttress is to prevent relaxation of the rock, it should be founded on a clean, sound rock surface. If this surface is not at right angles to the direction of thrust, then the buttress should be anchored to the base using steel pins to pre- vent sliding. Also, the top of the buttress should be poured so that it is in contact with the under- side of the overhang. In order to meet this second requirement, it may be necessary to place the last pour through a hole drilled downward into the cavity from the rock face, and to use a non-shrink agent in the mix. 12.4.6 Drainage As shown in Table 12.1, ground water in rock slopes is often a primary or contributory cause of instability, and a reduction in water pressures usually improves stability. This improvement can be quantified using the design procedures dis- cussed in Chapters 6–10. Methods of controlling water pressure include limiting surface infiltra- tion, and drilling horizontal drain holes or driving adits at the toe of the slope to create outlets for the water (Figure 12.16). The selection of the most appropriate method for the site will depend on such factors as the intensity of the rainfall or snow melt, the permeability of the rock and the dimensions of the slope. Surface infiltration. In climates that experi- ence intense rainfall that can rapidly saturate the slope and cause surface erosion, it is beneficial for stability to construct drains both behind the crest and on benches on the face to intercept the water (Government of Hong Kong, 2000). These drains are lined with masonry or concrete to prevent the collected water from infiltrating the slope, and are dimensioned to carry the expected peak design flows (see Figure 1.1(a)). The drains are also interconnected so that the water is dis- charged to the storm drain system or nearby water courses. Where the drains are on steep gradients, it is sometimes necessary to incorporate energy dissipation protrusions in the base of the drain to limit flow velocities. In climates with high rain- fall there is usually rapid vegetation growth, and periodic maintenance will be required to keep the drains clear. Horizontal drain holes. An effective means of reducing the water pressure in many rock slopes is to drill a series of drain holes (inclined upwards at about 5 ◦ ) into the face. Since most of the ground water is contained in discontinuities, the holes should be aligned so that they intersect the dis- continuities that are carrying the water. For the conditions shown in Figure 12.4, the drain holes are drilled at a shallow angle to intersect the more persistent discontinuities that dip out of the face. If the holes were drilled at a steeper angle, parallel to these discontinuities, then the drainage would be less effective. There are no widely used formulae from which to calculate the required spacing of drill holes, but as a guideline, holes are usually drilled on a spacing of about 3–10 m, to a depth of about one- half to one-third of the slope height. The holes are often lined with perforated casing, with the perforations sized to minimize infiltration of fines that are washed from fracture infillings. Another aspect of the design of drain holes is the disposal Stabilization of rock slopes 305 Lined collector drain Slope immediately behind crest graded to prevent pools of surface water from gathering during heavy rain Lined surface drain to collect run-off before it can enter top of tension crack Vertical pumped drainage well Horizontal hole to tap base of tension crack Potential tension crac k Potential slide surface Sub-surface drainage gallery Collector drain Horizontal hole to drain potential slide surface Fan of drill holes to increase drainage efficiency of sub- surface gallery Figure 12.16 Slope drainage methods. of the seepage water. If this water is allowed to infiltrate the toe of the slope, it may result in degradation of low-strength materials, or pro- duce additional stability problems downstream of the drains. Depending on site conditions, it may be necessary to collect all the seepage water in a manifold and dispose of it at some distance from the slope. Drain holes can be drilled to depths of sev- eral hundred meters, sometimes using drilling equipment that installs the perforated casing as the drill advances to prevent caving. Also, it is common to drill a fan of holes from a single set up to minimize drill moves (Cedergren, 1989). Drainage adits. For large slides, it may not be possible to reduce significantly the water pressure in the slope with relatively small drain holes. In these circumstances, a drainage tunnel may be driven into the toe of the slide from which a series of drain holes are drilled up into the saturated 306 Stabilization of rock slopes rock. For example, the Downie Slide in British Columbia has an area of about 7 km 2 and a thick- ness of about 250 m. Stability of the slope was of concern when the toe was flooded by the con- struction of a dam. A series of drainage tunnels with a total length of 2.5 km were driven at an elevation just above the high water level of the reservoir. From these tunnels, a total of 13,500 m of drain holes was drilled to reduce the ground water pressures within the slope. These drain- age measures have been effective in reducing the water level in the slide by as much as 120 m, and reducing the rate of movement from 10 mm/year to about 2 mm/year (Forster, 1986). In a mining application, ground water control measures for the Chuquicamata pit in Chile include a 1200 m long drainage adit in the south wall, and a num- ber of pumped wells (Flores and Karzulovic, 2000). Methods of estimating the influence of a drainage tunnel on ground water in a slope include empirical procedures (Heuer, 1995), theoretical models of ground water flow in homogeneous rock (Goodman et al., 1965), and three-dimensional numerical mod- eling (McDonald and Harbaugh, 1988). In all cases, the flow and drawdown values will be estimates because of the complex and uncer- tain relationship between ground water flow and structural geology, and the difficulty of obtaining representative permeability values. Empirical procedures for calculating inflow quantities are based on actual flow rates measured in tunnels. Based on these data, a relationship has been developed between the normalized steady- state inflow intensity (l/min/m tunnel length/m head) and the rock mass conductivity determined from packer tests (Heuer, 1995). The flow quant- ities can be calculated for both vertical recharge where the tunnel passes under an aquifer, and radial flow for a tunnel in an infinite rock mass. This empirical relationship has been developed because it has been found the actual flows can be one-eighth of the calculated theoretical values based on measured conductivities. Approximate inflow quantities can also be estimated by modeling the drainage adit as an infinitely long tunnel in a homogeneous, isotropic porous medium, with the pressure head on the surface of the tunnel assumed to be atmospheric. If flow occurs under steady-state conditions such that there is no drainage of the slope and the head above the tunnel H 0 is constant with time, the approximate rate of ground water flow Q 0 per unit length of tunnel is given Q 0 = 2πKH 0 2.3 log(2H 0 /r) (12.11) where r is the radius of the tunnel driven in homo- geneous material with hydraulic conductivity K. For rock formations with low porosity and low specific storage it is likely that transient condi- tions will develop where the head diminishes with time as the slope drains. An important aspect of slope drainage is to install piezometers to monitor the effect of drain- age measures on the water pressure in the slope. For example, one drain hole with a high flow may only be draining a small, permeable zone in the slope and monitoring may show that more holes would be required to lower the water table throughout the slope. Conversely, in low permeability rock, monitoring may show that a small seepage quantity that evaporates as it reaches the surface is sufficient to reduce the water pressure and significantly improve stability conditions. 12.4.7 “Shot-in-place” buttress On landslides where the slide surface is a well- defined geological feature such as a continuous bedding surface, stabilization may be achieved by blasting this surface to produce a “shot-in-place” buttress (Aycock, 1981; Moore, 1986). The effect of the blasting is to disturb the rock surface and effectively increase its roughness, which increases the total friction angle. If the total friction angle is greater than the dip of the slide surface, then sliding may be halted. Fracturing and dilation of the rock may also help reduce water pressures on the slide surface. Stabilization of rock slopes 307 The method of blasting involves drilling a pat- tern of holes through the slide surface and placing an explosive charge at this level that is just suffi- cient to break the rock. This technique requires that the drilling begins while it is still safe for the drills to access the slope, and before the rock becomes too broken for the drills to operate. Obviously, this stabilization technique should be used with a great deal of caution because of the potential for exacerbating stability con- ditions, and probably should only be used in emergency situation when there are no suitable alternatives. 12.5 Stabilization by rock removal Stabilization of rock slopes can be accomplished by the removal of potentially unstable rock; Figure 12.17 illustrates typical removal methods including • resloping zones of unstable rock; • trim blasting of overhangs; • scaling of individual blocks of rock. This section describes these methods, and the cir- cumstances where removal should and should not be used. In general, rock removal is a preferred method of stabilization because the work will eliminate the hazard, and no future maintenance will be required. However, removal should only be used where it is certain that the new face will be stable, and there is no risk of undermining the upper part of the slope. Area 4 on Figure 12.17 is an example of where rock removal should be carried out with care. It would be safe to remove Weathered rock Fresh rock 1. 2. 3. 4. Resloping of unstable weathered material in upper part of slope Access bench at top of cut Removal of rock overhang by trim blasting Removal of trees with roots growing in cracks Hand scaling of loose blocks in shattered rock Figure 12.17 Rock removal methods for slope stabilization (TRB, 1996). [...]... behavior of rock falls as they roll and bounce down slope faces (Piteau, 1 980 ; Wu, 1 984 ; Descoeudres and Zimmerman, 1 987 ; Spang, 1 987 ; Hungr and Evans, 1 988 ; Pfeiffer and Bowen, 1 989 ; Pfeiffer et al., 1990; Azzoni and de Freitas, 1995) Figure 12.19 shows an example of the output from the rock fall simulation program RocFall (Rocscience, 2004) The cross-section shows the trajectories of 20 rock falls,... 316 Stabilization of rock slopes 12.6.4 Rock catch fences and attenuators During the 1 980 s, various fences and nets suitable for installation on steep rock faces, in ditches and on talus run-out zones were developed and thoroughly tested (Smith and Duffy, 1990; Barrett and White, 1991; Duffy and Haller, 1993) Nets are also being used in open pit mines for rock fall control (Brawner and Kalejta, 2002)... pit and hillside showing geology and extent of slope failure; (b) regressive slope movement when mining at 184 0 m level; (c) slope movement over 13 months leading to slope failure; (d) slope velocity over two months prior to failure (Wyllie and Munn, 1979) Movement monitoring in the overturned siltstone beds at the crest of the pit When mining on the 184 0 m bench, a series of cracks formed on the 186 0... benched slopes is in tropical areas with deeply weathered rock and intense periods of rain In these conditions, lined drainage ditches on each bench and down the slope face are essential to collect runoff and prevent scour and erosion of the weak rock (Government of Hong Kong, 2000) Stabilization of rock slopes (a) 311 953 Elevation (m) 946 939 932 925 24 30 36 42 48 54 Off-set (m) 60 66 72 78 Height... height and face angle of the slope; a ditch design chart developed from field tests is shown in Figure 12.21 (Ritchie, 1963) The figure shows the effect of slope angle on the path that rock falls tend to follow, and how this influences ditch design For slopes steeper than about 75◦ , the rocks tend to stay close to the face and land near the toe of the slope For slope angles between about 55◦ and 75◦... 12.6 Protection measures against rock falls An effective method of minimizing the hazard of rock falls is to let the falls occur and control the distance and direction in which they travel Methods of rock fall control and protection of facilities at the toe of the slope include catchment ditches and barriers, wire mesh fences, mesh hung on the face of the slope and rock sheds A common feature of all... toe and the roll-out distance The report includes design charts that show, for all combinations of slope geometry, the relationship between the percent rock retained and the width of the catchment area Stabilization of rock slopes Slope gradient Free fall 0.25:1 0.3:1 0.5:1 0.75:1 1:1 1.25:1 40 0.1 120 7.62 m 2.44 m Slope height H (feet) 100 30 2.13 m 180 6.1 m 20 160 140 falling blocks, the slope. .. occurs on the fabric layers, and that they can withstand high energy impacts without significant damage (Barrett and White, 1991) Also, a 4 m high geofabric and soil barriersuccessfully withstood impact from boulders with volumes of up to 13 m3 and impact energies of 5000 kJ on Niijima Island in Japan (Protec Engineering, 2002), and a similar 1 .8 m wide geofabric barrier stopped rock impacts delivering 950...3 08 Stabilization of rock slopes the outermost loose rock, provided that the fracturing was caused by blasting and only extended to a shallow depth However, if the rock mass is deeply fractured, continued scaling will soon develop a cavity that will undermine the upper part of the slope Removal of loose rock on the face of a slope is not effective where the rock is highly degradable,... 75◦ , falling rocks tend to bounce and spin with the result that they can land a considerable distance from the base; consequently, a wide ditch is required For slope Figure 12.20 Configuration of benched cut in horizontally bedded shale and sandstone, with weaker coal and shale formations located at toe of cut faces angles between about 40◦ and 55◦ , rocks will tend to roll down the face and into the . they roll and bounce down slope faces (Piteau, 1 980 ; Wu, 1 984 ; Descoeudres and Zimmerman, 1 987 ; Spang, 1 987 ; Hungr and Evans, 1 988 ; Pfeiffer and Bowen, 1 989 ; Pfeiffer et al., 1990; Azzoni and de. in cracks Hand scaling of loose blocks in shattered rock Figure 12.17 Rock removal methods for slope stabilization (TRB, 1996). 3 08 Stabilization of rock slopes the outermost loose rock, provided. bench and down the slope face are essential to collect runoff and prevent scour and erosion of the weak rock (Government of Hong Kong, 2000). Stabilization of rock slopes 311 24 42 6030 48 6636