Wall Control by
Lyall Workman Calder & Workman Inc. 2501 Twin City Dr. Suite 2 Mandan, ND 58554 Tel. (701) 667-5785 Fax (701) 667-5784 e-mail:
[email protected]
WALL CONTROL BLASTING
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TABLE OF CONTENTS
1.0 Introduction................................................................
1
1.1 Definition of Controlled Blasting........................
1
1.2 The Importance of Controlled Blasting...............
1
1.3 Methods in Use....................................................
3
2.0 General Principles.....................................................
5
2.1 Introduction........................................................
5
2.2 Controlling the Energy Input and the Borehole Pressure..............................................................
6
2.2.1 Fully Coupled Borehole Pressure..............
8
2.2.2 Decoupling and Decking...........................
11
2.3 The Buffer Row....................................................
15
2.4 Effect of Water on a Decoupled Explosive Charge. 17 3.0 Influence of Conditions at the Site............................
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3.1 Principle Rock Properties.....................................
19
4.0 Wall Control Practices in Surface Operations...........
23
4.1 Explanation of Methods.......................................
23
4.1.1 Buffer Blasting...........................................
23
4.1.2 Presplitting.................................................
28
4.1.2.1 General Discussion......................
28
4.1.2.2 Spacing Between Holes...............
29
4.1.2.3 Presplitting on an Angle...............
31
4.1.2.4 Choosing the Hole Diameter........
34
4.1.2.5 Shooting the Presplit Line...........
37
4.1.2.6 Active highwall Presplitting in Dragline Operations..................... 4.1.3 Cushion Blasting........................................
40 45
4.1.4 Line Drilling................................................
55
4.1.5 Air Deck-Air Shock Techniques.................
56
4.2 Blast Design for Final Wall Shots........................
60
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5.0 Wall Control Practice Underground...........................
63
6.0 Controlled Blasting on Construction Projects...........
75
References.........................................................................
81
Appendix A: Technical Papers on Control Blasting.........
Considerations in pre-split Blasting for Mines and Quarries by J. Lyall Workman and Peter N. Calder. Control Blasting at Sherman Mine by Peter J. Calder and John N. Tuomi. Considerations for Small Versus Large Diameter Presplit Blasting by J. Lyall Workman and Peter N. Calder.
A Method for Calculating the Weight of Charge to use in Large Hole Presplitting for Cast Blasting Operations by J. Lyall Workman and Peter N. Calder. Wall Control Blasting at the Manassas Quarry by J. Lyall Workman and Peter N. Calder. Glossery Conversion Factors from Rock Slope Engineering by Hoek and Bray
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CONTROLLED BLASTING J. Lyall Workman and Peter N. Calder
1.0 INTRODUCTION 1.1 DEFINITION OF CONTROLLED BLASTING Controlled blasting refers to various techniques used to minimize damage to the rock at the limits of an excavation due to the action of the ground shock wave and the high pressure explosion gases, generated during the blast. 1.2 THE IMPORTANCE OF CONTROLLED BLASTING Wall control blasting techniques have been used in surface and underground blasting in the mining, quarrying and construction industries for many years. The specific reasons for the use of controlled blasting techniques may vary according to the industry and project, however, two generally applicable reasons can be identified.
1. To insure that the rock is broken to the excavation limit but not beyond. 2. To insure the subsequent safety of personnel and equipment, working under the wall, by avoiding backbreak and loose rock
on the face.
In open pit operations breakage beyond the pit limit is costly. Excessive backbreak at the perimeter generally results in an overall pit wall angle less than designed, and may result in the need for costly artificial support techniques. In fact, failure to properly
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3 control blasting at the final pit wall can cost a large open pit mine many millions of dollars in additional waste removal for the same ore mined (Workman and Calder, 1992). Underground, overbreak in the stope results in costly ore dilution. Poor breakage control at the perimeter of drifts and shafts means more scaling of the walls and roof and more difficulty installing support and facilities. In construction blasting breakage beyond the designed limits may lead to the removal of many tons of rock not specified in the contract. Added scaling and support may be needed for the long term stability of the wall. The consumption of concrete and other construction items may well increase. All of this is expensive. Equally important as cost, in every industry, is the need to provide a safe working environment. Pit and quarry walls that have sustained substantial backbreak are prone to hazardous rock falls. Safety benches, intended to arrest the fall of loose material will typically be narrow and ineffective. Drifts and stopes experiencing excessive overbreak will be more prone to hazardous rock falls. Similar hazards will also exist in construction work as well. Therefore, any organization that emphasizes safety will want to control blasting at the limits of an excavation.
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4 1.3 METHODS IN USE There are four principal controlled blasting techniques which are: • • • •
Presplitting Cushion blasting Buffer blasting Line drilling
Presplitting is the most commonly used technique especially in surface work. This is followed by cushion blasting, also known as trim blasting in open pits. Smooth blasting, used underground, is similar to cushion blasting. Buffer blasting may be used alone in cases where the rock is quite competent, but this is not a common approach. However, a properly designed buffer row at the back of the final production shot is essential to the success of most presplitting and cushion blasting applications. Line drilling involves the drilling of closely spaced small diameter holes at the perimeter of the excavation. These holes are not loaded with explosive, but form a discontinuity at the excavation limit. This method is costly because of the many boreholes drilled and is therefore only seen in blasting for civil works projects, where backbreak can be a very expensive result. Modified forms of line drilling may be used in mining and quarrying in special circumstances.
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5 2.0 GENERAL PRINCIPLES 2.1 INTRODUCTION Direct damage to the excavation limit due to blasting is usually found in the form of backbreak or overbreak, crest fracture and loose rock on the face. The mine operator has a number of tools available for minimizing or eliminating these problems. Techniques include changing the explosive type, or changing the blasthole diameter, by decoupling the explosive, by decking, and by changing the burden and spacing. Changing the depth of subgrade drilling or the stemming height can reduce crest fracture and any resultant narrowing of the width of safety benches. Changing the millisecond delay timing and the rotation of the round may also be helpful in eliminating these problems. The rock characteristics and geology must be considered when designing controlled blasts as these have an important influence on the final results. The compressive strength, crushing strength and tensile strength of the rock should be known. The frequency and orientation of joints and fractures in the rock are also important parameters. These variables cannot be controlled but must be determined by suitable field and laboratory techniques. Geology can have pronounced effects on the results of wall control blasts. For example, it is known that trim blasting does not work well in the presence of relatively shallow dipping joint planes dipping into the
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6 excavation (Crosby and Bauer, 1982). It may not always be possible to obtain the classic result, with the half-barrel of all the wall control holes showing on the face, when adverse geology is encountered. However, if backbreak, crest fracture and face loose rock have been minimized, then the result will be far more acceptable than a wall in the same rock where no controlled blasting has been performed. This can be clearly seen in figure 1, where the upper bench has been presplit while the lower one has not. Furthermore, there is evidence to indicate that good results can be obtained, even when the ground is heavily fractured or the rock is very weak (Workman and Calder, 1993, 1992). 2.2 CONTROLLING THE ENERGY INPUT AND THE BOREHOLE PRESSURES A fundamental goal of all wall control blasting is to reduce the energy input and the borehole pressures at the perimeter of the excavation. The borehole pressures generated by commercial explosives, that are fully coupled to the hole, are much greater than the rock strength and will cause extensive damage around the blasthole. Therefore, these pressures must be reduced.
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Figure 1: Pit Wall Illustrating the Difference Between a Presplit Bench (Upper) and a Bench with no Wall Control Blasting at the Perimeter (Lower)
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8 2.2.1 FULLY COUPLED BOREHOLE PRESSURE The borehole pressure for a fully coupled hole can often be obtained from the manufacturer of the product being considered for use. However. in the absence of this information it can be calculated using the following formula: (P b ) c = NqD 2 where (Pb = Borehole pressure of a fully coupled charge completely filling the blasthole q = Specific gravity (density) of the explosive in gm/cc D = Velocity of detonation of the explosive confined in a fully coupled blasthole of the given diameter N = Constant determined from figure 2 or 3 depending on the units being used While this equation may not yield exact results it has proven quite adequate for practical design requirements. However, it cannot be used in the case of aluminized explosives. The velocity of detonation is reduced because the initial reactions of the oxidizer with aluminum are endothermic. However, beyond the detonation zone the equilibrium shifts to the very rapid formation of exothermic reaction products. Therefore, the actual borehole pressure will be considerably higher than that calculated from the detonation velocity. Low density explosives produce low borehole pressures because the detonation velocity is reduced. Table 1 lists borehole pressures for ANFO charges of different density. Low density mixes were made with microballoons or perlite (Calder. 1977).
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Figure 2: Chart for Determining N Given the Specific Gravity of the Explosive (Imperial Units)
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11 2.2.2 DECOUPLING AND DECKING A primary means of reducing the borehole pressure is to decouple the charge from the hole. This means that the diameter of the charge is less than the diameter of the hole. Pressure may be further reduced by decking, whereby wooden or cardboard spacers are used between charges or the charges are taped to detonating cord with a gap left between individual cartridges. The net coupling ratio can be expressed by: d
C.R. = C % d hc where C = the percent of explosive column actually loaded dc = charge diameter dh = hole diameter For a given hole diameter and explosive the usual approach is to decouple radially first. if this is insufficient to reduce the borehole pressure enough than decking can be employed.
Table 1: Borehole Pressure Generated by ANFO at Different Densities. ANFO Density gms/cc
Detonation Detonation Borehole Borehole Velocity Velocity Pressure Pressure ft/sec m/sec psi MPa
0.80
13500
4116
364125
2511
0.40
9200
2805
84553
583
0.30
8200
2500
50378
347
0.25
7000
2134
30593
211
0.20
6600
2012
21758
150
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12 When a charge is decoupled from the blasthole the explosion gases
must expand to fill the hole volume before exerting borehole pressure. Therefore the decoupled borehole pressure will be much less than the coupled value. The decoupled pressure may be calculated from the following formula:
(P b ) dc = (P b ) c % (C.R. ) 2.4 where (Pb)dc = The borehole pressure for a decoupled and/or decked charge C.R. = Coupling ratio Figure 4 is a graph of the coupling ratio versus the coupling ratio to the 2.4 power. If one is known the other can be found. In using these equations it is necessary to have an idea of what an acceptable decoupled borehole pressure will be. In presplitting it has been found that the pressure should be in the range of 2 to 5 times the uniaxial compressive strength. (Calder and Tuomi, 1980) The upper bound is the crushing strength which should not be exceeded. In larger hole diameters it is often better to set the decoupled borehole pressure near to the uniaxial compressive strength of the rock because of the greater radius of rupture that may result around larger diameter boreholes, when the borehole pressure exceeds thecompressive strength of the rock. This potential for large rupture radius around the borehole can lead to a wall more prone to unravel over time.
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15 In the case of cushion blasting the coupling ratio should not exceed 0.45. While the borehole pressures generated in cushion blasting are higher than those employed in presplitting, these must be considerably reduced from the fully coupled values for good results. For purposes of illustration figure 5 shows the decoupled borehole pressure for 3-inch charges of ANFO in various hole diameters. In this case there is no decking and all reduction in the pressure is obtained from radial decoupling. In some presplitting applications a concentrated charge is used in or near the bottom of the hole with the remainder of the borehole left void. Upon detonation the explosion gases are free to expand up the hole and exert a suitable decoupled pressure on the surrounding rock. This method has been used extensively in active highwall presplitting when blast casting in dragline mines. It has also been used in other types of mining, generally being most successful if the ground is reasonably competent thereby avoiding damage at the bottom of the hole and excessive leakage of gases as these expand up the borehole. 2.3 THE BUFFER ROW Occasionally buffer blasting alone may be sufficient to protect a final excavation limit from damage. However, when presplitting or cushion blasting the last row of the final production blast must be a buffer row. The exceptions to this rule would be when active highwall presplitting for a
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16 dragline operation or in small diameter work underground where a buffer row is not always used. The buffer row must be designed with a sufficient charge to break the rock between the buffer hole and the final wall. However, the explosive consumption in the buffer row must not be so great as to cause breakage beyond the plane of the final wall or the controlled blasting effort will have been wasted. Often, when damage is observed beyond the final wall limit the problem is the buffer row design rather than the presplit or trim row. The buffer row is designed with less explosive in the hole than is found in production blasting boreholes. Because the explosive is kept low, in the hole, with a greater length of stemming above, there is less potential for crest fracture and face loose rock. but the toe between the buffer hole and the excavation limit can still be adequately broken. The low center of gravity of the charge in the buffer hole causes it to behave like a spherical charge, for which cube root scaling applies (Livingston, 1957). In a buffer row a scaled depth of burial (SDOB) of about 1.5 times the optimum scaled depth of burial for the given explosive in the given rock type should be used. The scaled depth of burial is simply the depth from the surface to the center of the charge column divided by the cube root of the total explosive weight in the column. Ideally the charge should have a length not exceeding 8 times the diameter of the borehole. If, because of the hole depth or diameter, the charge length exceeds 8
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17 times the diameter the calculation should be performed using the depth to the center of a charge column equal in length to 8 times the diameter and located at the top of the charge. Only the explosive weight contained in this charge, at the top of the column, should be used in the calculation. The depth to the center of the charge can be calculated as follows: C = SDOB x W 1/3 where D = Distance from the upper bench surface to the center of gravity of the top eight diameters of the charge SDOE = Scaled depth of burial W 1/3 = Cube root of the weight of explosive found in a column length equal to 8 times the diameter As an example in hard jointed rock a scaled depth of burial of 4.0 ft/lb1/3 (1.59 m/kg1/3) has often been found suitable. Table 2 shows the optimum scaled depth of burial and the recommended first approximation values for the scaled depth of burial in a buffer row. These values must be taken as general guidelines only, for not every possibility of jointing, rock type and subgrade drilling can be accounted for.
Also, different
explosives in the same rock may yield different optimum scaled depths of burial. Therefore, field optimization is usually required. 2.4 EFFECT OF WATER ON A DECOUPLED EXPLOSIVE CHARGE When a decoupled charge is surrounded by water the pressure generated by the detonating explosive, at the borehole wall, will be considerably higher than would be the case if the explosion gases were free to expand across an air filled gap.
The degree of decoupling
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18 achieved will be much less than that calculated assuming the charge is surrounded by air.
In fact because water is quite incompressible the
pressure transferred to the borehole wall may be quite similar to that of a fully coupled hole. The explosive charge will need to be decoupled to a greater extent than normal. If the area can be dewatered prior to final wall blasting this will be the best solution. it will be necessary to choose a fully waterproof explosive for this application. When a column of water exits above a concentrated presplit charge at the bottom of a large diameter hole another problem can develop. The water column tends to behave as stemming and the explosion gases are inhibited from freely expanding up the hole. There will be more damage around the bottom of the hole. The presplit crack may not extend the full length of the borehole. These holes will work best if pumped before explosive loading. They should be loaded and fired promptly to minimize the water column that forms above the explosive charge.
3.0 INFLUENCE OF CONDITIONS AT THE SITE The properties of the rock and the site geology are of significant importance when designing a controlled blast. If these factors are ignored the results will be, at best, a hit and miss affair. Serious backbreak. crest fracture, face loose rock or sliding of weak portions of the wall are all possible outcomes.
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19 It is also important to recognize that in complex geological settings it may not be possible to achieve the classic result. However, even though the half-casts of all the holes are not visible on the face the controlled blast will still have been successful if a safe, stable wall has been achieved at an economical cost. Table 2: First Approximation Scaled Depth of Burial at the Collar of the Buffer Row Holes
Rock Type
Range of Optimum SDOB, ft/lb1/3
Range of SDOB for Use on Buffer Row ft/lb1/3
Very hard massive
2.2—2.5
3.30—3.75
Hard more fractured
2.5—3.0
3.75—4.50
Medium
3.0—3.5
4.50—5.25
Soft
3.5—4.0
5.25—6.00
Very Soft
4.0—4.5
6.00—6.75
3.1 PRINCIPLE ROCK PROPERTIES The most important rock properties are the tensile strength, compressive strength and crushing strength. Also very important are the nature, frequency and orientation of joints and fractures, the rock density, longitudinal wave velocity and Young's Modulus. Ideally these properties should be measured in-situ. In-situ values reflect the effects of weathering and structural features in the rock. A rock which tests as quite strong in the laboratory may be considerably weaker when weathering, groundwater alteration, presence of structures such as open joints, bedding or
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20 foliation planes and fractures due to previous blasting are accounted for. However, at this time methods for measuring rock properties in-situ are not very satisfactory and are usually costly. Therefore, laboratory tests are generally relied on. Laboratory data can be adjusted by a site factor to account for in-situ conditions. Deciding what the site factor should be is not a simple task and will be an approximation. Most practical is to design the controlled blast based on the laboratory results and observe the results in the field. Then the design can be adjusted to account for any problems until an optimum result is obtained. It may then be possible to back calculate the in-situ uniaxial compressive strength and tensile strength. Backbreak and radial crushing around the borehole result when the stress produced in the rock by the explosion exceeds the crushing strength of the rock. The crushing strength is typically two to five times the uniaxial compressive strength. Major backbreak problems are likely if an explosive loading that was successful in competent ground is subsequently used in highly jointed or fractured ground, even though the rock type is the same. Therefore, powder factors and decoupled borehole pressures must be adjusted to account for structural conditions and the actual crushing strength of the rock surrounding the hole. The potential for wall damage due to structural features is less when the joints are tight or infilled and possess some strength. When the
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21 joints are open and have little strength the potential for backbreak and crest fracture is much greater. The orientation of the joints has a major influence on the controlled blast results. When joints or fractures strike parallel to the excavation face a smooth clear wall may be obtained. When the joints are steeply dipping (>70°) the wall can be made to conform to the joint planes. When the joints are more shallow dipping it is undesirable to cause the wall angle to conform to these planes. There is greater chance that planes will undercut the face. When this occurs it is more difficult to obtain a classic result because there is a greater likelihood that portions of the wall will slide off along these structured planes. Large diameter cushion blasting has been found unsuited to these conditions. Presplitting may be more successful if great care is taken to design the presplit and buffer rows to minimize the disruption experienced on the joint planes. It takes relatively little movement along the plane to destroy cohesional resistance and cause the material resting on the joint to be more prone to slide. When steeply dipping joints dip back into the wall while striking parallel to the face, sliding on undercut planes is not possible. However, toppling failures may occur. In the presence of these features the final wall should not be vertical. An angle of 70 to 80 degrees is more suitable. A toe buttressing effect is provided and the wall is far more likely
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22 to remain safe and in good condition for the long term (Workman and Calder, 1992). When structural features strike at angles other than parallel to the face the amount of backbreak depends on the nature of the joints and fractures and their strike. Open joints are likely to break back more than tight, infilled joints. Planes striking at 45 degrees to the face are likely to break back further than near vertical joints striking at 90°. The frequency of jointing is important. Jointing begins to interfere with wall control results when the joint spacing is less than the hole spacing. In presplitting the hole spacing should not exceed twice the major joint spacing. Frequent jointing can lead to greater crest fracture. The explosive collar height must be increased or the upper column load reduced. When the stress due to the reflected ground shock wave at the free face, near to a blast, exceeds the rock tensile strength slabbing can occur. If joints, bedding planes or foliations exist, striking parallel to the face, the potential for slabbing is greatly increased. Slabbing is especially a hazard when blasting near to tunnels or when blasting in a pit that is in close proximity to the walls of another pit. Reduced explosive loading may be necessary. Where rock breakage is not desired, as in the case at the final excavation limit, rock properties that relate to the in-situ rock strength are important. The Young's Modulus of Elasticity is a measure of the
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23 brittleness of a rock and its susceptibility to backbreak. Rock with a high Young's Modulus has a higher crushing strength and is harder to break. Higher borehole pressures may be permissible at the perimeter. Rocks with a higher longitudinal wave velocity are also usually found to be stronger. Weaker rock or strata that has been weakened by weathering, alteration or fracturing due to dense jointing or previous blasting exhibits a lower longitudinal wave velocity. This fact leads to the seismic techniques for determining overburden depth, depth of broken rock, radius of rupture, jointing and density. As an in-situ method these techniques may be particularly valuable for determining the nature of the in-place rock.
4.0 WALL CONTROL PRACTICES IN SURFACE OPERATIONS 4.1 EXPLANATION OF METHODS 4.1.1 BUFFER BLASTING This is perhaps, the simplest form of wall control shooting. The last row of the production blasting pattern is altered to limit the energy input at
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24 the final wall. The explosive loading is reduced and as a consequence the burden and spacing are also decreased. As described in section 2 explosive loading is often reduced by selecting a scaled depth of burial greater than would normally be used. Another approach is to use decoupled bagged powder above a toe load of fully coupled explosive. Buffer blasting can only be used as the sole controlled blasting technique when the ground is quite competent. Some minor backbreak or crest fracture may develop but this will be much less than would be caused by the production blast holes. Where buffer blasting can be used alone the cost of wall control will be quite economical. Figure 6 illustrates a typical buffer blast design. In most cases buffer blasting is used in conjunction with another wall control blasting technique. A properly designed buffer row is very important to most successful presplit or trim blasts. Design of the buffer row is the same as when the technique is used alone. It becomes important to
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Figure 6: A Typical Buffer Blast Design
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26 insure that the buffer row is at the correct location relative to the presplit or trim row. Typical design for the buffer row includes using a scaled depth of burial at the top of the charge of 1.5 times the production hole value and reducing the powder factor to 0.5 - 0.8 times the production row powder factor. Burdens range from 0.5 to 0.75 times the production burden. The spacing should not be less than the burden and will usually be 1.0 to 1.25 times the buffer row burden. To avoid backbreak and crest fracture the buffer row holes must be properly located in front of the intended plane of the final wall or the presplit line. This distance must be sufficiently large to insure that the stress at the final wall is adequately attenuated to avoid crushing beyond the plane of the wall. Figure 7 shows how the stress generated by detonating buffer row holes attenuates with distance from the blasthole. From this chart one can see that in quite soft rock, such as coal mine overburden, spacing the buffer row 10 feet or more in front of the presplit line may indeed be prudent. In hard rock the spacing at the toe needs to be much less to break the rock between the buffer row and the presplit line. However, breakage beyond the presplit can be avoided. This chart also shows that, to avoid crest fracture in competent rock, drilling the presplit holes on an angle is advantageous. One can space the presplit and buffer hole closely at the toe for breakage while obtaining a greater standoff at the crest. When the compressive and crushing where
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28 strength of the rock are known figure 7 can be helpful in determining to place the buffer row relative to the presplit line. At the same time the buffer row should not be moved out too much or poorly fragmented material may be frozen to the wall and a toe may be left at the base of the intended face. In hard rock it has been found that the toe of the buffer row should be 3 to 5 feet (1 to 1.5 meters) from an intended face angled at 80 degrees. In soft rock, such as coal overburden, it has been necessary to move the toe of the buffer row out as much as 15 feet (4.6 meters) to keep the zone of crushed material from extending beyond the planned wall location. 4.1.2 PRESPLITTING 4.1.2.1 General Discussion Presplitting is the most common controlled blasting technique and has proven successful in applications from large open pit mines to civil construction. This method involves the drilling of closely spaced holes at the planned excavation perimeter which are lightly loaded with explosive in order to generate an appropriate borehole pressure as described in previous sections. Presplitting is being done using hole diameters ranging from 2 inches to 12¼ inches. Often, small diameter presplitting is preferred for technical reasons and because the cost per square foot of wall may be lower (Calder and Tuomi, 1980; Workman and Calder, 1989).
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Other
29 mines use large diameter holes in order to employ the same drills for presplitting as for production drilling. This approach has worked especially well in active highwall presplitting designs associated with blast casting operations. It has not always been as successful in other types of mining applications. In small diameters (6 inches, 152 mm) hole spacings of 5 to 18 feet (1.5-5.5M) have been employed. As spacings become larger geological structure becomes an increasingly important control on this dimension. 4.1.2.2 Spacing Between Holes The spacing between the holes is a function of the hole diameter, decoupled borehole pressure and the tensile strength of the rock. The tangential stress is expressed as: r 2 T = (P b ) dc % rh2 T
= tangential stress
(Pb)dc = decoupled borehole pressure rh
= radius of the borehole
r
= the distance from the center of the hole to the point of measurement
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30 It can be shown that the stress extending between two boreholes fired together is: T = 2(Pb )dcrh This stress must everywhere exceed the force resisting crack formation which is related to the hole spacing and the tensile strength. This leads to the spacing equation, which is: (Sanden, 1974) S =
dh
P b dc +T
12T
where S = spacing between presplit holes, ft. T = rock tensile strength psi dh = hole diameter, inches For the radius in inches, the decoupled borehole pressure in psi and the tensile strength in psi. the spacing is given in inches. For appropriate metric units it will be in centimeters. This formula points out the importance of knowing the tensile strength of the rock (measured using the Brazilian Test) in order to properly compute the spacings. Table 3 gives first order approximations of tensile strength for typical materials. The spacing between presplit holes may have to be varied in different areas of the pit if differing rock types exist with different uniaxial compressive strengths and tensile strengths.
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31 Table 3: First Order Approximations of Tensile Strength for Different Rock Types Tensile psi
Strength MPa
Rock Type
Example
Hard
Granite, Taconite
1,600—6,0 11.03—41. 00 37
Limestone
800—1,600 5.52—11.0 3
Medium Low
Asbestos Ore, Coal Overburden
