Chapter 34: Cast Blasting

This chapter addresses the topic of cast blasting for dragline surface coal mining operations. Surface coal mining uses large draglines to remove the overburden and expose the coal seams for extraction. Cast blasting is a blasting technique designed to use explosive energy to move or "cast" a significant portion of the overburden material into the previously mined cut, reducing the amount of material a dragline must rehandle. In cast blasting the explosive accomplishes two tasks: (1) fragmentation of the in-place overburden and (2) displacement of this material into the mine pit.

Cast blasting is a proven technique providing economic benefits for surface coal operations that use draglines to strip overburden. The benefits of cast blasting include:

• Increased stripping production
• Reduced dragline rehandling
• Reduced operating costs
• Increased equipment life
• Reduced fuel consumption

Cast blasting concepts have evolved from the first attempts in the 1970s. Cast blasting has become increasingly popular as a means of improving dragline productivity.

BLASTING AND DRAGLINE PERFORMANCE

Dragline productivity for a given overburden depth will depend on (1) operating factors, (2) equipment factors, and (3) the blasting program. The blasting program affects dragline productivity through its impact on the fragmentation of the muckpile, the looseness of the muckpile, and the amount of material cast into the pit.

Fragmentation

Overburden fragmentation must be suitable for efficient dragline digging. Oversized material increases cycle time and decreases bucket fill. Proper fragmentation is achieved by correctly matching the explosive energy distribution to the rock mass characteristics. Draglines work most efficiently when the fragmented rock is small enough to flow into and fill the bucket without hang-ups or bridging. Uniform fragmentation throughout the muckpile is also important since inconsistent fragmentation leads to variable cycle times.

Muckpile Looseness

Muckpile looseness or swell affects the ease of digging. A tight, poorly fragmented muckpile is difficult to dig and increases cycle time and wear on the dragline components. Conversely, a loose muckpile allows the bucket to easily penetrate and fill. Cast blasting should produce sufficient swell to facilitate efficient digging while avoiding excessive throw that scatters material beyond the intended cast distance.

Cast Distance

The further material is cast into the pit, the less material the dragline must rehandle. Cast distance depends on several factors including explosive energy, timing, burden, bench height, and geology. The percentage of material cast over the highwall varies considerably depending on site conditions and blast design. Typical cast percentages range from 20% to 45% of the total overburden volume.

Figure 34.1 illustrates the relationship between dragline productivity and overburden depth with and without cast blasting. As shown, cast blasting can significantly increase productivity, especially as overburden depth increases.

SOURCES OF POOR BLASTING RESULTS

Poor cast blasting results can be attributed to many factors. These factors generally fall into six categories: (1) geologic factors, (2) drilling factors, (3) explosives factors, (4) initiation system factors, (5) design factors, and (6) pre-blast factors. Understanding these factors is essential for troubleshooting and improving cast blast performance.

Geologic Factors

Poor Overburden Fragmentation

Poor overburden fragmentation can result from massive, competent rock that resists breakage. The presence of hard layers or massive sandstone beds that do not break readily can cause oversized material that reduces digging efficiency. Blasts in these formations may require increased powder factors or modified explosive products to achieve adequate fragmentation.

Loose Muckpile Issues

A muckpile that is too loose, characterized by excessive voids and low density, can cause problems for dragline operations. The bucket may not fill properly as material falls away during the digging cycle. This condition often results from geological conditions such as weak, fractured rock or coal seams that break into excessively small fragments.

Back Row Issues

Back row problems occur when the material at the back of the blast does not break and move properly. This can result in a tight area that is difficult for the dragline to dig. Back row issues often relate to insufficient energy at the back of the pattern or geology that does not respond well to the applied explosive energy.

Inadequate Cast

Inadequate cast occurs when insufficient material is thrown into the pit. This increases the dragline's workload and reduces productivity. Causes include insufficient explosive energy, improper timing, excessive burden, and unfavorable geological conditions such as water-saturated material that dampens the explosive effect.

Inconsistent Material Movement

Inconsistent material movement across the blast creates an uneven muckpile that is difficult to dig efficiently. The dragline operator must constantly adjust to varying conditions, increasing cycle times and reducing productivity. This problem often results from geological variations across the blast pattern that cause different responses to the explosive energy.

Inappropriate Geometry

Inappropriate geometry refers to conditions where the pit configuration does not allow for efficient cast blasting. Factors such as highwall angle, pit width, and bench height affect cast blast performance. When these factors are not properly considered in the blast design, poor casting results can occur.

Geologic Anomalies

Geologic anomalies such as faults, clay seams, weathered zones, and water-bearing strata can significantly affect cast blast performance. These features can cause:

• Energy absorption that reduces cast distance
• Preferential breakage along weakness planes
• Uncontrolled material movement
• Poor fragmentation in certain zones

Parting Failures

Parting failures occur when weak horizontal layers within the overburden allow the upper portion of the bench to separate from the lower portion. When this happens, explosive energy is vented through the parting rather than being used to fragment and cast the overburden. Parting failures can significantly reduce cast efficiency and may cause flyrock hazards.

Drilling Factors

Drilling factors that affect cast blast performance include:

• Hole deviation: Crooked holes change burden and spacing relationships, affecting energy distribution
• Incorrect hole depth: Holes that are too shallow or too deep affect break at the toe and energy distribution
• Incorrect pattern layout: Errors in burden and spacing reduce blast efficiency
• Drilling in wrong location: GPS errors or surveying mistakes can place holes in incorrect positions

Explosives Factors

Explosives factors affecting cast blast performance include:

• Incorrect product selection: Using an explosive with insufficient energy or water resistance
• Product quality issues: Degraded or improperly mixed products may not perform to specifications
• Incorrect loading: Improper loading procedures can leave gaps or create overloaded sections
• Sleep time: Extended sleep times can allow water infiltration or product degradation

Initiation System Factors

Initiation system factors that affect cast blast performance include:

• Timing errors: Incorrect delay times can cause poor fragmentation and reduced cast
• Detonator failures: Misfires or out-of-sequence detonations disrupt the blast progression
• Improper detonator placement: Incorrect primer placement affects explosive column initiation

Excessive Misfires

Excessive misfires create serious problems in cast blasting operations. Misfires can result from:

• Water damage to explosives or detonators
• Cutoffs from rock movement during the blast
• Faulty detonators or initiation systems
• Improper loading procedures

Misfires must be carefully managed following established safety procedures and regulations.

Pre-Blast Problems

Pre-blast problems that can affect cast blast performance include:

• Wet holes: Water in boreholes can degrade explosives or prevent proper loading
• Caving holes: Unstable boreholes that collapse before or during loading
• Blocked holes: Obstructions that prevent proper loading depth
• Loading errors: Incorrect charge weights or deck configurations

CAST BLAST DESIGN

The goal of cast blast design is to optimize the distribution of explosive energy to achieve maximum displacement and muckpile looseness required. High face velocities, relative to other blasting applications, are necessary, but massive energy not used is likely to be lost in non-casting applications, even if this is achieved at the expense of some of the shock energy that might be available with other products (See chapter 9).

Caution Explosives with sufficient energy must be chosen for the toe burden.

When selecting explosives for cast blasting, heave is an important parameter to achieve the displacement and muckpile looseness required. High face velocities, relative to other blasting applications, are necessary. Therefore, explosives products that maximize heave energy are likely to be best in most casting applications, even if this is achieved at the expense of some of the shock energy that might be available with other products (See chapter 9).

Caution Excessive toe distances for the explosive used will inhibit cast blast performance.

Excessive toe distance will cause the overburden to topple around a fixed pivot point at the toe, reducing the cast. Coal damage increases because the energy, unable to move the toe, is directed toward the least resistance, which often is the coal seams. Therefore, properly matching the energy input to the toe distance is very important to good casting performance and to minimizing coal damage.

Burden Adjustments

Burden adjustments, to achieve similar results at the toe, can be made for an alternate explosive based on the burden and relative bulk strength ratio that demonstrates known performance (See equation 34.1).

$$R = \frac{B_k}{RBS_k^{1/3}}$$ <!-- VERIFIED -->

Equation 34.1

Where:

• R = Ratio of burden to relative bulk strength of charge of known performance
• B_k = Toe burden of known explosive performance (meters/feet)
• RBS_k = Relative bulk strength of explosive of known performance (dimensionless)

Once this ratio is known for any explosive, the toe burden for an alternate explosive can be calculated using equation 34.2.

$$B = R \times RBS_E^{1/3}$$ <!-- VERIFIED -->

Equation 34.2

Where:

• B = Toe burden for alternate explosive (meters/feet)
• R = Ratio of burden to relative bulk strength of charge of known performance
• RBS_E = Relative bulk strength of alternate explosive

EXAMPLE 34.1

Calculate the toe burden for an emulsion/ANFO blend with an RBS_E of 1.2 when the mine has successfully used ANFO with a RBS of 1.00 to pull an 8 meter toe burden on the front row.

Step 1 Calculate the burden to RBS ratio using equation 34.1 for the explosive of known performance (ANFO).

$$R = \frac{B_k}{RBS_k^{1/3}} = \frac{8}{1^{1/3}} = \frac{8}{1}$$ <!-- VERIFIED -->

R = 8

The burden to relative bulk strength ratio of the known explosive is 8.

Step 2 Calculate the toe burden for the ANFO/emulsion blend with RBS_E of 1.2 using equation 34.2.

$$B = R \times RBS_E^{1/3} = 8 \times 1.2^{1/3}$$ <!-- VERIFIED -->

$$B = 8 \times 1.063$$ <!-- VERIFIED -->

B = 8.5

The toe burden is 8.5 meters.

EXAMPLE 34.2

Calculate the toe burden for an emulsion/ANFO blend with an RBS_E of 1.2 when the mine has successfully used ANFO with a RBS of 1.00 to pull a 26 foot toe burden on the front row.

Step 1 First calculate the burden to RBS ratio for the explosive of known performance (ANFO).

$$R = \frac{B_k}{RBS_k^{1/3}} = \frac{26}{1^{1/3}}$$ <!-- VERIFIED -->

R = 26

The Burden to relative bulk strength ratio of the known explosive is 26.

Step 2 Now calculate the toe burden for the ANFO/emulsion blend with RBS_E of 1.2.

$$B = R \times RBS_E^{1/3} = 26 \times 1.2^{1/3}$$ <!-- VERIFIED -->

$$B = 26 \times 1.063$$ <!-- VERIFIED -->

B = 27.6

The toe burden is 27.6 feet.

Caution In soft overburden the toe distance may change by an amount approaching the change in the relative bulk strength.

Therefore, calculations can begin using the cube root of the RBS and expand as appropriate to optimize.

Please note that in soft overburdens the change in toe distance may factor by an amount approaching the direct value of the relative bulk strength. Therefore, one can start at the cube root and expand as appropriate to optimize.

Explosives with the same bulk strength will provide similar results on the same patterns when shot in the same rock. However, results may not be identical. The resulting fragmentation and blast profile will depend to some degree on how the energy is partitioned. The shock energy (brisance) is always less than the heave energy (See chapter 11). However some explosives partition more energy into shock while other classes of products have proportionally more heave energy.

Stemming Length

Proper energy distribution within the blast is very important to successful cast blasting. Achieving proper distribution directly depends on explosive selection, pattern layout, and the delay timing sequence. The relationship to explosive selection discussed above ensured the use of energy-producing products in the borehole.

A method for calculating the stemming height so that as boreholes are involved examines the energy release of the top of the charge. A cratering approach (See chapters 14 and 9) is used to determine the length of stemming that will provide good breakage without uncontrolled flyrock. Cast blast design includes locating the explosive collar to allow good horizontal displacement of the crest area moved by the top of the bench. Therefore, scaled depth of burial (SDOB) may be somewhat less in a given rock and geologic setting than would be the case for other blasting applications. The control of flyrock due to the cratering effect at the borehole collar as discussed in chapter 15.

As a rule, to produce good top fragmentation without excessive throw, the collar height in a massive brittle rock can be estimated from the scaled depth of burial of the top of the charge equal to 1.1 meters/kilogram^(1/3) (2.8 feet/pound^(1/3)). For softer formations which are typical in coal mine overburdens, this would increase to values approaching 1.4 meters/kilogram^(1/3) (3.5 feet/pound^(1/3)) to 4.0 feet/pound^(1/3)). The latter case is shown in figure 34.7 which illustrates cratering results in sandstone (after Livingston, 1956). The same situation would also be true for finely laminated predominantly shale. The best way to optimize the stemming height is to do a high-speed camera study. Ideally this film shows the point at which the best throw is obtained without uncontrolled flyrock resulting.

Once the correct stemming height is chosen and adequate energy is supplied at the toe the next important requirement is the powder factor and the associated burden and spacing. The need for maximum displacement requires a higher powder factor than is used in conventional blasting for draglines, where fragmentation is the main goal. Powder factors of 0.59 kilograms/meter³ to 0.75 kilograms/meter³ (1.0 pound/yard³ to 1.25 pounds/yard³) ANFO equivalent are commonly used in cast blasting, whereas powder factors in the range of 0.26 kilograms/meter³ to 0.45 kilograms/meter³ (0.35 pound/yard³ to 0.75 pound/yard³) ANFO equivalent are more common when fragmentation is the only concern. Thus, an increase in powder factor in the range of 2 to 2.5 times the normal explosives consumption is common. These are examples of commonly used ratios only. For a specific situation, experimentation is necessary to determine the optimum product at the optimum load.

The increase in energy can be achieved by using a more energetic explosive, decreasing the burden and spacing or a combination of the two. In cast blasting the overburden drilling costs range from $2.54/meter to $4.91/meter ($0.88/foot to $1.50/foot). This is so low that the economics may favor denser patterns rather than more energetic products. Strata explosive costs can vary over time as a result of ingredient pricing, so it is good practice to check the economics periodically.

Delay Orientation

Three patterns are commonly used for cast blasting: (1) square, (2) staggered square, and (3) equilateral (See chapter 14). These are chosen for their regular arrangement to enhance energy distribution and the application of delay timings for maximum displacement.

Delay Patterns

Patterns can be delayed in different ways. The V-1 and V-2 delay patterns result in effective burdens less than the drilled burden (See figures 34.11 and 34.9 respectively). This effect is used to advantage in achieving displacement, an important requirement in cast blasting.

Figure 34.9 plots the ratio of center radius to scaled burden in hard rock. As the scaled burden is decreased below 1.36 meters/kilogram^(1/3) (3.0 feet/pound^(1/3)), which is the optimum scaled depth of burial in this rock tested, the crater radius continues to increase while the depth of the crater becomes more shallow. The crater assumes more of a bowl shape. The radius may be as much as twice the crater depth. (See chapters 14 and 15.) Fragmentation becomes finer. Displacement increases as the scaled burden decreases. Although coal overburden may be weaker than the rock tested and the optimums SDOB smaller, the trend remains the same.

The curve in figure 34.8 shows that reduction of the burden to a value lower than the full crater value can be accompanied by an increase in upward heave. This is important because burden reduction through the delay pattern is accompanied by an increase in effective spacing. These cratering results show that certainly for the delay patterns in figures V-1 and V-2 use configuration of the increased effective spacing can be broken out successfully.

This effect is often achieved in practice by drilling in a square pattern and then delaying it in such a manner as to produce a blast "en-echelon" or along the diagonals. This gives a burden to spacing ratio of 1:2, without the disadvantage that a rectangular row-by-row blast would produce with a large spacing along the front. This is a V-1 delay pattern and while it reduces the burden, the 45° angle to the highwall (the principle free face) is usually too sharp to maximize castover.

A square pattern can be used or it can be modified to the staggered square pattern. (See figure 34.9). Another option is the equilateral triangle pattern where the drilled spacing is greater than the drilled burden by a factor of 1.15 (See chapter 14).

It is quite common to delay the cast blast in a row-on-row method directly toward the open pit. This is especially useful where an irregular highwall requires fill-in boreholes on the front row to control toe distances.

An alternative that can be considered is the staggered equilateral pattern. This configuration provides the best distribution of energy throughout the blast. This is followed closely by the staggered square pattern. However, the equilateral triangle pattern can be delayed at a flatter angle to the principal free face (highwall) when the V-2 delay pattern is used, which is an advantage. This pattern is best used with a straight highwall and where high casting percentage drops the overburden substantially leaving a partial free face along the side. The blast can then be oriented across the full and partial free face. The V-2 delay pattern as shown in figure 34.9 is useful in starting at the end of a new pit where there is no partial free face to the side of the blast. This would then be modified to the staggered equilateral pattern shot on the V-2 arrangement as is indicated in figure 34.10 for the remaining blasts in the dragline pit.

Good success has been achieved using the staggered equilateral pattern (See figure 34.10) in cast blasting for reasons related to both the energy distribution and delay options. Therefore, those designing cast blasts should consider this pattern type.

In general, the square pattern is not recommended for casting when delayed row-on-row (See figure 34.13), since it has the poorest energy distribution. When delayed on the V-1 delay pattern (See figure 34.11), the echelon angle of 45° is too steep to create the displacement required in figures 34.1 and 34.2 (See Castover at beginning of this chapter).

Blasts designed to throw overburden into the previous strip mine cut have been tied in with a variety of orientations. All have advantages and disadvantages that must be weighed. Figure 34.12 shows the V-3 pattern, where the point of opening is four boreholes with the remainder of the blast initiated on echelons at low angle to the free face to promote forward rock movement.

Caution In general, the V-3 delay pattern creates effective spacings that may not be efficient when casting.

If it is contemplated that the blast will be delayed row-on-row, a staggered square pattern is the best option. The advantages of good energy distribution are combined with the avoidance of wide spacing on the front row, which occur when a rectangular pattern is used. (See Figure 34.13). This approach orients movement directly at the principal free face, which is the current highwall. The rows of boreholes move successively into the mined pit. If there are delays within rows some boreholes in the next row may initiate before the previous row is completely detonated. However those boreholes lag far behind the previous row boreholes and the orientation of movement is essentially at right angles to the axis of the drilled rows.

When the blast is delayed row-on-row the echelon burden is the same as the drilled burden. There is no burden reduction available by the delay orientation.

Caution Only way to reduce the echelon burden in a row-on-row pattern is to decrease the drilled burden and increase the spacing to maintain the same area of influence per borehole.

The disadvantage to this practice is that spacings may become too large to obtain all the burden reduction that would be beneficial if the staggered square shown in figure 34.13 is modified to a rectangular pattern.

When the boreholes on the succeeding row lag those on the previous row by a smaller amount the detonation of the boreholes begins to take on a diagonal initiation sequence. This is accompanied by a reduction in effective burden compared to drilled burden. Common diagonal or echelon initiation configurations include the V-1, V-2, and V-3 patterns. The geometric characteristics of these are described in chapter 14.

There is an advantage to blasting with delay patterns that produce an effective burden smaller than the effective spacing (See chapter 14). The impact on fragmentation may be greater in brittle rock, but in either case free rock velocities and displacement will be enhanced.

A staggered equilateral pattern tied V-2 can be quite useful in casting shots. This arrangement gives good energy distribution and the greatest burden reduction. In casting it also provides the flatter angle to the principal free face (See V-2 configuration).

When delaying a staggered square or staggered equilateral pattern on the V-2 diagonal, the back row boreholes are pulled away from the new highwall one at a time (See figures 34.10 and 34.5 for examples of this effect). This gives a poking action, which is less damaging and avoids throwing material back against the active wall peopl. Figure 34.14 is an example of an echelon sequence on a diagonal equilateral delay sequence used in the cast block.

Effective Burden and Spacing

Reduced burden-spacing ratio is achieved with proper pattern geometry and delay application.

The burden reduction achieved increases as the pattern geometry changes from square to staggered square to staggered equilateral.

For example, a 270 millimeter (10 5/8 inch) diameter borehole on a 9.1 meter x 9.1 meter (30 foot x 30 foot) square pattern has an equivalent equilateral pattern of 8.5 meters x 9.8 meters (28 feet x 32 feet).

If the explosive in this example is ANFO at a 0.85 gram/centimeter³ density, the following blast design information is true. For the square pattern delayed on the V-1, the effective burden is 6.5 meters (21.3 feet) and the square root SDOB off the side of the cylindrical charge is 0.93 meters/kilogram^(1/3) (2.3 feet/pound^(1/3)).

For the staggered square arrangement delayed on the V-2 pattern, the effective burden is reduced to 5.1 meters (16.7 feet) and the SDOB is 0.73 meters/kilogram^(1/3) (1.94 feet/pound^(1/3)).

For the V-2 equilateral pattern illustrated in figure 34.14 the effective burden is 4.8 meters (16 feet) and SDOB is 0.70 meters/kilogram^(1/3) (2.5 feet/pound^(1/3)).

Effective Burden Reduction

Effective burden is reduced from the row-on-row square pattern using the staggered square and staggered equilateral patterns.

Staggered equilateral patterns require two conditions for successful application in cast blasting: (1) A high total castover is needed so that the face along the side of the blast next to the previous blast is partially free. In efficient casting programs the partial free face may be about sixty percent of the total overburden height. Then blasted material can move efficiently on a diagonal. (2) A straight highwall. With the staggered equilateral pattern it is somewhat more difficult to adjusted front row boreholes to accommodate an irregular face. Pattern layout limiting factors and requirements are summarized in table 34.6.

Patterns Layout Limiting Factors and Requirements

Table 34.6 – Pattern layout limiting factors and requirements.

Millisecond Delay Timing

It is important to provide an adequate time interval between successive series of boreholes detonating in a cast blast so that there is burden relief on one row of boreholes prior to the next period firing. The time it takes for the burden to start to move as a blast is proportional to the size of the burden. Put another way, it takes time for the fracturing process followed by the gas expansion to take place. If the delay time between boreholes firing on successive delays is not longer than the minimum detachment time, the blast can be choked and the digging hard because the formation remains tight and excessive coal seam damage may well occur. This requirement applies to delay times between rows or diagonals, not to delays within a row.

Caution Insufficient delay time intervals for successive delays must exceed the minimum rock detachment time to prevent tight digging and coal damage.

Figure 34.15 shows, from field experimentation, the minimum time at bench blasts that it takes before various burden distances start to move. These values were determined from high-speed movies of blasts with boreholes ranging from 165 millimeter to 305 millimeters (6½ inch x 12 inch) diameter. The graph shows that there is a delay time before the face starts to move after which it is displaced at a velocity characteristic of the burden involved.

Figure 34.16 plots the minimum time before face movement as a function of the burden. The relationship is reasonably a straight line showing the minimum time for relief to be 1 millisecond per foot of effective burden (3.3 milliseconds/meter).

Optimum times are longer than 3.3 milliseconds/meter (1 millisecond/foot) of effective burden. In normal non-cast blasting, 6.6 milliseconds/meter to 8.2 milliseconds/meter (2 milliseconds/foot to 2.5 milliseconds/foot) of effective burden is often suitable for achieving good relief, but actual optimums are site and rock specific.

Delay Time Optimization Experimentation at the mine is necessary to optimize the delay time.

The high-speed video camera is an important tool for this purpose.

Excellent relief is required to cast as much of the total overburden into the previous cut as possible. Coal overburden usually consists of rock with low compressive strength that is laminated and precracked. The strata absorb considerable energy before relieving. Long delay time intervals are often an effective strategy to allow for this energy absorption. For example, at one mine with overburden of 34.5 megapascals (5,000 pounds/inch²) compressive strength, optimum results occur when the delay time is 19.7 milliseconds/meter (6 milliseconds/foot) of effective burden or 11.8 milliseconds/meter (3.6 milliseconds/foot) of back-burden.

Optimum delay time intervals in casting can be expected to range from 9.8 milliseconds/meter to 19.7 milliseconds/meter (3 milliseconds/foot to 6 milliseconds/foot) of effective burden. Shorter time intervals apply to more competent, brittle strata. Rock lithology and structural geology differ between mines and site-specific optimization is recommended.

Caution Site-specific optimization is recommended, because lithological and structural geology differ between mines.

In some cases it proves advantageous to increase the delay interval between successive rows further back in the blast. This accounts for the fact that later rows must displace farther to clear the current cut and require increasing relief to achieve this result. It may be appropriate to increase the delay interval between rows by up to 100 milliseconds. This is especially the case when firing the blast row-on-row.

Limit surface delay scatter = result in cutoffs and misfires. Therefore, in-hole delays become almost mandatory for casting blasts. The in-hole delay should be long enough to ensure surface tie in advance of borehole detonation.

An alternative approach is to use all in-hole delays using different periods to obtain the required shot orientation. The only surface delays are an occasional bridge delay to restart the sequence once the highest in-hole period has been used. The advantage to this method is that surface initiation energizes boreholes ahead of the rock movement minimizing chance of cutoffs or misfires. The disadvantage is that some flexibility to change the tie-in to cater for field differences from design is lost once in-hole delays are placed.

Electronic delays are also available. As described elsewhere, these delays are highly accurate. Casting blasts depend on excellent relief for successive rows of boreholes to move out into the previously mined pit. One important requirement to achieve this is delay time precision and accuracy (See chapter 13). Electronic delay times are the most precise and accurate available, so it makes sense to consider timing when cast blasting. Programmable electronic delays also allow for in-hole initiation without the need for prepackage surface delay detonation.

Importance of Optimum Placement Front Row Boreholes

The location of the front row is very important in cast blasting. Good casting typically requires face velocities of 20 meters/second (65 feet/second) up to 30 meters/second (100 feet/second). When face velocities exceed 30 meters/second (100 feet/second), uncontrolled rock movement becomes a concern.

High face velocities will be hard to achieve if the front row is located too far from the face. Conversely, if front row boreholes are too close to the face gas venting will occur and performance will drop. An irregular highwall makes front row borehole placement difficult with fill-in boreholes likely to be needed. Face velocities and casting performance will suffer. Therefore, a straight, consistent highwall is needed for best results.

The high-speed video camera is an excellent tool for studying the effects of front row location. The face velocity can be determined from these videos. Gas venting can be seen if it occurs. The use of videos helps establish the optimum location.

Best placement of the front row is achieved if the highwall is profiled (See chapter 32) before the front row is designed. Targetless laser survey equipment is an excellent tool for this application and the recently developed photogrammetric system has also provided good results.

Once the optimum scaled burden is found for front row boreholes it should be kept constant. Changes in overburden type may require different positioning, and a change in explosive can affect the optimum scaled depth of burial and front row position as well.

The typical optimum SDOB will range between 0.62 meters/kilogram^(1/3) to 1.00 meter/kilogram^(1/3) (2.5 feet/pound^(1/3) and 4.0 feet/pound^(1/3)). Note that this is the square root scaling off the side of the long cylindrical column charge not the cube root scaling associated with breakage at the top or bottom of the charge. Higher optimum SDOB is associated with softer rock and is the more typical design criteria in coal mine overburdens. Figure 34.17 illustrates this relationship.

In some cases, it is necessary to restrict the amount of cast to achieve the needed muckpile profile. For example, sometimes the dragline elevation must be kept above the stacking height capability of the dragline. One way to restrict movement is by locating the front row further back. The greater scaled depth of burial will reduce the displacement of the overburden.

Caution The front row placement must have proper burden to create the desired muckpile profile.

It is very important that the toe distance on the front row be balanced with the available energy of the explosive in use (See Energy Input and Distribution section earlier in this chapter). Draglines frequently strip overburdens 30.5 meters (100 feet) or more in depth. Reduced highwall angles can lead to long toe distances. For particularly shallow angles, it can be difficult for any explosive to adequately fragment and move the toe (See figure 34.18).

For example, a 30.5 meter (100 foot) highwall at a 75° angle has a 8.2 meter (27 foot) offset from crest to toe (See chapter 18). When the drill is positioned 2.4 meters (8 feet) behind the crest, the toe distance is 10.7 meters (35 feet). ANFO would be near its performance limit in soft rock and would not be effective in harder materials except possibly in the larger borehole diameters. Other explosives could be used to obtain a satisfactory result at the toe.

If the same highwall is at an angle of 55°, the offset increases to 21.3 meters (70 feet). The 2.4 meter (8 foot) drill set back results in a 23.8 meter (78 feet) toe distance. Even the most energetic explosive will perform poorly in this case. Figure 34.18 is a graphical representation of these values. It includes the 2.4 meter (8 foot) drill setback. The setback may vary at different mines and the total distance at the graphic may need to be adjusted accordingly. For example, if the operating value is 3 meter setback (10 feet) then the toe distance will need to be increased by 1.8 meters (5 feet).

Caution Proper toe distance and energy factor at the toe are fundamental to cast blast design.

Long toe distances on the front row substantially reduce the castover percentage and lead to a poorly fragmented toe. In addition severe damage to the coal seam can result. Cause occurs where the coal seams has been impacted leaving a 3.1 meter to 4.6 meter (10 feet to 15 feet) wide trench in the coal in-filled with overburden. Even a parting below an upper seam can show on weak bedding planes. These phenomena occur when very long toe distances exceed relief of the toe and the energy (which is substantial given the high powder factors) is directed down into the coal seam.

Active highwall presplitting (discussed later in this chapter) is a technique often used to control the highwall angle. Angle drilling the production boreholes is also a very important method in many cases. Steeper highwall angles are better for cast blasting performance.

Caution The highwall angle must not exceed an angle required for proper ground control supported by a geotechnical study.

Relationship Of Blast Width To Highwall Height

The relationship between the height of the highwall and the pit width is very important to successful casting.

Caution When the pit width is too great for the highwall height, casting effectiveness is reduced.

As the pit becomes wider relative to the highwall height it is increasingly difficult for the material to clear the cut. Figure 34.20 illustrates the relationship between the total castover and the pit height-to-width ratio. This chart is based on total castover; any cast is relative to total. Clearly ratios approaching 1.0 give the best results. Because many factors affect the castover percentage the data is scattered and an upper and lower limit line are plotted as being the boundaries of possible castover results for a given height to depth ratio.

Therefore, mines adopting the casting technique must take the blasting needs into account when deciding on the pit width. It is no longer just a matter of how wide a pit the dragline can physically strip, or how wide the pit should be for convenient coal loading and hauling operations. Very wide pits will likely be inappropriate. Given overburden depths typically experienced in North American coal mines pit widths of 38 meters to 61 meters (125 feet to 200 feet) generally represent the limits of multiple pit width.

For mines in shallow overburden (less than 21 meters / 70 feet) strip casting may not be a useful technique. For example, if the overburden is 15.3 meters (50 feet) deep it will be difficult to design pit geometry wide enough for efficient coal loading while at the same time creating the geometry needed to cast into the prior cut properly. At 21.3 meters (75 feet) a 36 meter (125 feet) wide pit has a ratio of 0.60 that may permit reasonable casting results in a pit that is wide enough for efficient operation of coal loading and other procedures. Therefore in mines where pits should not be designed to do pits that both (1) maximize displacement and (2) obtain good fragmentation and loosening of the muckpile.

Caution Pits designed for cast blasting should be designed to both (1) maximize displacement and (2) obtain good fragmentation and loosening of the muckpile.

Relationship Of Blast Width To Blast Length

A blast is a three-dimensional event. Therefore, the length of the blast is also important to a successful cast blast. A short blast will have more corner effects (end effects) that inhibit the overburden displacement. By contrast these end effects are only a small percentage of the total effect for a long shot.

Based on the observations and evaluation of numerous cast blasts, the length of the blast should be 4 to 5 times its width (Workman, 1995). Therefore, in a 46 meter (150 feet) wide pit the blast length should be between 183 meters and 230 meters (600 feet and 750 feet) for best results.

Caution In theory, the longer the better but when the blasts become very long, with many boreholes, the chance for errors in tie-in and borehole misfires increases.

Blasts with many boreholes have an increased potential for dead pressing, sympathetic detonation, and detrimental effects of delay time variation. Therefore, while desiring to keep a good length to width ratio, it is also important to balance the safety and performance problems that occur from placing too many boreholes in a very large blast and attempting to tie in too many boreholes.

Figure 34.21 illustrates an actual blast pattern layout that has suitable dimensions for cast blasting. The dimensions on this drawing illustrate a specific cast blast and are not meant to be design guidelines since overburden varies. For V-2 equilateral pattern, the burden is 8.5 meters (27.9 feet) and the spacing is 9.8 meters (32.1 feet). Note that these are close to the recommended 4:1 to 5:1 length to width ratio. The same length to width dimensions applies if a row-on-row blast is used. The pit is 46 meters (150 feet) wide by 183.5 meters (116.5 feet) deep. It is 183 meters (600 feet) long, so this is about the preferred minimum width to length relationship. It is recognized that operational constraints may require shorter blasts periodically but this should be kept to a minimum. In this example 60% of the side wall is exposed.

Controlling the Castover

Usually cast blast design involves maximizing the castover. However, in other situations less than maximum cast may be desired. This can occur for example if the after blast profile from maximum casting would lower the dragline elevation to the point where the machine has inadequate stacking height.

There are several ways to reduce the amount of castover. Some combinations of these methods, suitable to the specific situation, would be chosen to restrict the motion. The techniques listed in table 34.9 are used to control the cast over.

Techniques To Control Castover

Table 34.9 – Techniques to control castover.

Where casting is desired, but less than the maximum amount, some combination of the above factors will yield the desired result.

ACTIVE HIGHWALL PRESPLITTING

Many mines using cast blasting incorporate active highwall presplitting into the operation. Each successive highwall is presplit to provide a regular highwall in excellent condition. Some operations also put the coal. The presplit may be vertical or incises with competent rock having little rock structure or at an angle of dipping joints or joint planes on curve followed by the coal. This is called a half pipe wall and has esthetic advantages for cast blasting.

The difference between the presplit line and the back row of the production blast is referred to as the new highwall. These boreholes may be initiated before the production blast is drilled or tied in with the production blast, but timed to fire in a manner leading the progression of the production shot.

Unlike the technique discussed in chapter 36, a highly loaded buffer row or in cast used it is necessary to adequately fragment the overburden next to the new highwall. Therefore, it is important to optimum the distance between the presplit line and the back row of the production blast to achieve adequate breakage while avoiding damage into the new wall.

In addition, when the casting blast is delayed along the diagonal echelons, the back row of production boreholes are stripped away from the new highwall one at a time. This reduces the amount of energy being driven back against the presplit wall at any moment in time and helps to avoid damage to the new highwall.

It is important to remember that the U.S. Mine Safety and Health Administration (MSHA) requires a ground control plan that includes safe highwalls, so casting procedures must work within the ground control portion.

A very important characteristic of a well-presplit face is that accurate control on the burden for the full depth on the front face row boreholes becomes much easier to realize (See figures 34.22 and 34.23). This enhances control of the cast blast (Crosby, W.A. and Bauer, A., 1982). The front row can be located to give the desired SDOB off the side of the front row blast holes. High face velocities can be achieved. The distances in the figures are for illustration only and mines must determine the best configurations for their specific site. Also, in many cases the wall will be "angle" presplit to avoid stability problems.

For operations with ground water problems, it may be possible to use presplitting to dewater the boreholes. To presplit for water control, the sides of the block out to the existing highwall are presplit in addition to the new highwall. The block is detached from the surrounding mass and drains. For success the block needs to have sufficient mass permeability to drain readily.

This can enable the use of non-waterproof explosives rather than waterproof products. Depending on the costs of the explosives used and the number of wet boreholes this presplitting to dewater the rock mass may provide a cost benefit.

Caution When coal seams have significant dip toward the face (empty pit) the unblasted rock in front of the presplit row may shift too much when the presplit row is initiated if the sides have also been presplit for dewatering.

For example, in an actual pit where the coal seams is flat lying and the perimeter of a 46 meter (150 feet) wide by 305 meter (1,000 feet) long pit is fully presplit, the entire presplit block may sway up to 50 centimeters (20 inches) toward the empty pit. The main boreholes drilled after the presplit is detonated are usually found to be completely dry.

Caution Lateral movement during presplitting is most common when the overburden bedding plane-coal seam contact is weak.

Keycut

The keycut is the slot produced adjacent to the new highwall by the dragline or other equipment at the initial step of digging each block of overburden to uncover the coal.

The use of the presplit line clearly defines the fragmented material for the drag line operator. Bulldozer assistance is often used in the keycut because the blasted and displaced overburden usually drops too much to allow a conventional dragline "keycut." The dragline does not need to dig back against the wall and a very clean face is achieved. Figures 34.22 and 34.23 show the contrast between a presplit and non-presplit section of highwall.

Presplit Line Design

This section discusses important factors in the design of an active highwall presplit. These include the factors listed in table 34.10.

In some cases other equipment, such as a backhoe, is used to clean the new highwall and the keycut is eliminated.

Design Factors For Active Wall Presplitting

Table 34.10 – Design factors for active wall presplitting.

The two design factors discussed here are the (1) presplit explosive charges and (2) spacing between boreholes.

AHP Charges

The horizontally bedded overburdens, combined with the pronounced plane of weakness at the overburden/coal interface allows presplitting to be achieved using a concentrated charge located at or near the toe of the hole. This is in contrast to presplitting at most open pits and quarries where a decoupled charge of explosives is distributed up the borehole to approximately 1.83 meters to 3.05 meters (6 feet to 10 feet) from the collar.

The borehole above the charge is left empty. Upon detonation the explosion gases are free to expand up the borehole, thereby reducing the borehole pressures to an acceptable level to create the presplit crack without causing wall damage around the boreholes.

Methods for calculating the diameter of a decoupled presplit charge are summarized here to complete the discussion. The discussion of presplit line design follows closely procedures found in Part 7 of the Crater B1 Slope Manual with the addition of calculating the weight of explosive in a concentrated charge at the bottom of the borehole equivalent to a decoupled charge distributed up the presplit borehole (Workman & Calder, 1991).

For decoupled charges, the first step is to calculate the explosion gas pressures for the fully coupled condition using equation 34.3.

$$P_b = N \times \rho_e \times c_d^2$$ <!-- VERIFIED -->

Equation 34.3

Where:

• P_b = Coupled borehole pressure (bars) (pounds/inch²)
• ρ_e = Explosive density (grams/centimeter³)
• c_d = Explosive detonation velocity (feet/second) (meters/second)
• N = Constant (See figure 34.24)

EXAMPLE 34.3

Calculate the coupled borehole pressure for a quarry that is presplitting using an explosive with a density of 1.25 grams/centimeter³ and a detonation velocity of 4,575 meters/second.

Step 1 Determine the metric value of N from figure 34.24.

N is 1.3 × 10⁻³ (0.00130)

Step 2 Calculate the coupled borehole pressure using equation 34.3

$$P_b = N \times \rho_e \times c_d^2$$ <!-- VERIFIED -->

$$P_b = 0.00130 \times 1.25 \times 4{,}575^2$$ <!-- VERIFIED -->

$$P_b = 0.00130 \times 1.25 \times 20{,}930{,}625$$ <!-- VERIFIED -->

$$P_b = 34{,}012$$ <!-- VERIFIED -->

The coupled borehole pressure is 34,012 bars.

EXAMPLE 34.4

Calculate the coupled borehole pressure for a quarry that is presplitting using an explosive with a density of 1.25 grams/centimeter³ and a detonation velocity of 15,000 feet/second.

Step 1 Determine the U.S. value of N from figure 34.24.

N is 1.78 × 10⁻³ (0.00178)

Step 2 Calculate the coupled borehole pressure using equation 34.3.

$$P_b = N \times \rho_e \times c_d^2$$ <!-- VERIFIED -->

$$P_b = 0.00178 \times 1.25 \times 15{,}000^2$$ <!-- VERIFIED -->

$$P_b = 0.00178 \times 1.25 \times 225{,}000{,}000$$ <!-- VERIFIED -->

$$P_b = 500{,}625$$ <!-- VERIFIED -->

The coupled borehole pressure is 500,625 pounds/inch².

The value of N varies with the density of the explosive. Figure 34.24 is a graph that relates the constant N to explosive density.

In many cases, it may be possible to obtain an accurate coupled borehole pressure from the manufacturer and avoid the necessity of calculations. Once the coupled borehole pressure has been obtained the decoupled pressure can be calculated using equation 34.4.

$$(P_b)_{dc} = P_b \times (R_c)^{2.4}$$ <!-- VERIFIED -->

Equation 34.4

Where:

• (P_b)_dc = Decoupled borehole pressure, bars/(pounds/inch²)
• R_c = Coupling ratio

The usual practice in active highwall presplitting, where decoupled explosives are used in a large diameter borehole, is to set the decoupled borehole pressure equal to a suitable multiple of the compressive strength of the rock in question. The coupling ratio to the 2.4 power is then the ratio of the decoupled to coupled pressure. The coupling ratio can be determined and, therefore, the ratio of charge to a given borehole diameter.

For cast blasting, the "coupling ratio" is the 2.4 power, which allows for the expansion of explosive gases to the borehole between the explosive and the borehole wall. Solving equation 34.4 for the "coupling ratio" produces equation 34.5.

Caution In coal mines, the pressure in large-diameter boreholes can exceed the compressive strength to a degree and still give a good result. A factor of 3 times the compressive strength seems to work well, which often amounts to about 103.4 megapascals (15,000 pounds/inch²) to 172.4 megapascals (25,000 pounds/inch²) of expanded borehole pressure. This is just an approximation. Field trials are required of the same to determine the best explosive amount. Good results will not cause damage. Beyond 10.4 kilograms (23 pounds) where blast loading on a nearby neighbor is considerable, other amounts may be appropriate.

$$R_c = \left(\frac{(P_b)_{dc}}{P_b}\right)^{1/2.4}$$ <!-- VERIFIED -->

Equation 34.5

EXAMPLE 34.5

Calculate the "coupling ratio" for the example 34.3 (34,012 bars), where the desired decoupled borehole pressure is 1,020 bars using equation 34.5.

$$R_c = \left(\frac{(P_b)_{dc}}{P_b}\right)^{1/2.4}$$ <!-- VERIFIED -->

$$R_c = \left(\frac{1{,}020}{34{,}012}\right)^{1/2.4}$$ <!-- VERIFIED -->

$$R_c = (0.03)^{1/2.4}$$ <!-- VERIFIED -->

$$R_c = 0.2319$$ <!-- VERIFIED -->

The coupling ratio is 0.2319.

EXAMPLE 34.6

Calculate the "coupling ratio" for the example 34.4 (500,625), where the desired decoupled borehole pressure is 15,000 pounds/inch² using equation 34.5.

$$R_c = \left(\frac{(P_b)_{dc}}{P_b}\right)^{1/2.4}$$ <!-- VERIFIED -->

$$R_c = \left(\frac{15{,}000}{500{,}625}\right)^{1/2.4}$$ <!-- VERIFIED -->

$$R_c = (0.03)^{1/2.4}$$ <!-- VERIFIED -->

$$R_c = 0.2319$$ <!-- VERIFIED -->

The coupling ratio is 0.2319.

$$R_c = C^{1/2} \times \left(\frac{d_e}{d}\right)$$ <!-- VERIFIED -->

Equation 34.6

Where:

• R_c = Coupling ratio (based on coupled and decoupled borehole pressures)
• C = Percent of column loaded* (length basis)
• d_e = Diameter of charge, millimeters** (inches)
• d = Diameter of borehole (millimeters) (inches)

*Accounts for decoupling along the axis by leaving a gap between individual charges. **The diameter of the explosive charge must be calculated to determine the weight of explosives to load.

Equation 34.7, derived from equation 34.6, is used to calculate the explosive charge diameter.

$$d_e = \frac{R_c \times d}{C^{1/2}}$$ <!-- VERIFIED -->

Equation 34.7

Where:

• d_e = Diameter of charge, millimeters (inches)
• R_c = Coupling ratio (based on coupled and decoupled borehole pressures)
• d = Diameter of borehole (millimeters) (inches)
• C = Percent of column loaded* (length basis)

Charge diameter provides for decoupling between the charge and charge diameters.

EXAMPLE 34.7

Calculate the decoupled explosive charge diameter for a 270 millimeter diameter borehole and a "coupling ratio" of 0.2319 using equation 34.7. Assume C = 1.

$$d_e = \frac{R_c \times d}{C^{1/2}}$$ <!-- VERIFIED -->

$$d_e = \frac{0.2319 \times 270}{1}$$ <!-- VERIFIED -->

$$d_e = 62.6$$ <!-- VERIFIED -->

The explosive diameter is 62.6 millimeters.

EXAMPLE 34.8

Calculate the decoupled explosive charge diameter for a 10.625 inch diameter borehole and a "coupling ratio" of 0.2319 using equation 34.7. Assume C = 1.

$$d_e = \frac{R_c \times d}{C^{1/2}}$$ <!-- VERIFIED -->

$$d_e = \frac{0.2319 \times 10.625}{1}$$ <!-- VERIFIED -->

$$d_e = 2.46$$ <!-- VERIFIED -->

The explosive diameter is 2.46 inches.

An important factor to determine is the desired magnitude of the decoupled borehole pressure. The borehole pressure must exceed the dynamic tensile strength of the rock by a significant amount. This is entirely possible since the tensile strength is often a factor of 8 to 10 less than the compressive strength. Unlike fracturing into the highwall next to presplit boreholes is to be avoided.

Caution The borehole pressure in active wall presplitting should not exceed the dynamic compressive strength of the rock by more than 3 times and in particularly weak overburden this may need to be decreased further.

To most accurately design an active highwall presplit for a specific mine overburden, core samples should be obtained and uniaxial compressive strength (UCS) tests should be performed to determine the U.S. value, a sufficient number of tests should be done to obtain a statistically valid result.

Active wall presplitting most commonly implies a concentrated fully coupled charge of bulk loaded or packaged explosive in the bottom of the borehole. Therefore, it is necessary to determine the weight of charge to be placed. The amount of charge is dependent on the diameter and depth of the presplit borehole, which would contain the volume of gases produced by the distributed decoupled column. A greater borehole volume will require an additional weight of explosive to maintain the same pressure after expansion because the volume is greater.

The decoupled charge diameter required for the presplit borehole as a continuous decoupled column is determined using equation 34.7 above. Then the weight required can be calculated for a given presplit depth with equations 34.8 (metric) and 34.9 (U.S.).

$$Q = 0.0785 \times d_e^2 \times l_c \times \rho_e$$ <!-- VERIFIED -->

Equation 34.8 (Metric)

$$Q = 0.3403 \times d_e^2 \times l_c \times \rho_e$$ <!-- VERIFIED -->

Equation 34.9 (U.S.)

Where:

• Q = Charge weight (kilograms) (pounds)
• d_e = Diameter of charge, centimeters (inches)
• l_c = length of borehole loaded if a distributed, decoupled charge were used (meters) (feet)
• ρ_e = density of explosive (grams/centimeter³)

The value of (l_c) is the borehole depth minus the length of borehole left unloaded at the top. This can often be set at 3.05 meters (10 feet) but may vary depending on the nature of the rock. The prescribed charge weight is then placed at the bottom of the borehole. It may be either bulk loaded or a loaded with a packaged product. Using packaged explosive since a small degree of decoupling naturally occurs around the charge, which may help to avoid or reduce damage to the borehole wall in the immediate vicinity of the charge.

EXAMPLE 34.9

Calculate the total weight of explosive (density = 1.25 grams/centimeter³) to be loaded as a concentrated charge in a presplit program. Boreholes are 36.5 meters deep, and the top of the explosive column is estimated to be 3.05 meters below the collar. The decoupled diameter of the explosive was determined to be 6.25 centimeters and the explosive density is 1.25 grams/centimeter³.

Step 1 Calculate the length of the decoupled charge as follows:

L = l_bottom − l_collar

L = 36.5 − 3.05

L = 33.45

The length of the decoupled charge is 33.45 meters.

Step 2 Calculate the weight of explosive to be loaded as a concentrated charge using equation 34.8.

$$Q = 0.0785 \times d_e^2 \times l_c \times \rho_e$$ <!-- VERIFIED -->

$$Q = 0.0785 \times 6.25^2 \times 33.45 \times 1.25$$ <!-- VERIFIED -->

$$Q = 128.1$$ <!-- VERIFIED -->

The weight of explosive to load as a concentrated charge is 128.1 kilograms.

This weight should then be loaded as a concentrated charge in the bottom of the borehole with the gases free to expand up the presplit borehole and decrease in pressure in order to create a presplit line, but not damage the new highwall.

EXAMPLE 34.10

Calculate the total weight of explosive (density = 1.25 grams/centimeter³) to be loaded as a concentrated charge in a presplit program. Boreholes are 120 feet deep and the top of the explosive column is estimated to be 10 feet below the collar. The decoupled diameter of the explosive was determined to be 2.46 inches and the explosive density is 1.25 grams/centimeter³.

Step 1 Calculate the length of the decoupled charge as follows:

L = l_bottom − l_collar

L = 120 − 10

L = 110

The length of the decoupled charge is 110 feet.

Step 2 Calculate the weight of explosive to be loaded as a concentrated charge using equation 34.9.

$$Q = 0.3403 \times d_e^2 \times l_c \times \rho_e$$ <!-- VERIFIED -->

$$Q = 0.3403 \times 2.46^2 \times 110 \times 1.25$$ <!-- VERIFIED -->

$$Q = 283$$ <!-- VERIFIED -->

The weight of explosive to load as a concentrated charge is 283 pounds.

This weight would then be loaded as a concentrated charge in the bottom of the borehole with the gases free to expand up the presplit borehole and decrease in pressure in order to create a presplit line but not damage the new highwall.

Two important factors in active wall presplitting are that (1) deeper boreholes require more weight of explosive to maintain the same borehole pressure after expansion because the volume is greater and (2) presplit boreholes of larger diameter require more explosive weight to sustain the same pressure after expansion because there is more volume.

This can be observed in table 34.10, which tabulates the weight of Heavy ANFO (HANFO) needed to generate approximately 103.4 megapascals (15,000 pounds/inch²) of expanded pressure in the borehole. It can be clearly seen that the required weight increases as the hole depth increases for a given hole diameter and the explosive weight increases as the hole diameter increases for a given presplit hole depth.

The HANFO used in this example table has a density of 1.26 grams/centimeter³ and a detonation velocity of 4,878 meters/second (16,000 feet/second). The charges are designed to produce a borehole pressure of 103.4 megapascals (15,000 pounds/inch²). This is a generic example. An operation will need to determine charge weights specific to that mine that the trends will be the same.

Weight Of HANFO Required To Generate 15,000 psi Of Decoupled Borehole Pressure

Table 34.10 – HANFO loading density required to generate 15,000 psi of decoupled borehole pressure. (Courtesy: L. Workman)

A study of table 34.10 shows the required explosive weight becomes substantial in deep boreholes. The increase in charge weight with depth is linear for the weight at 45.7 meters (150 feet) to be 50% greater than the weight at 30.5 meters (100 feet). Also note that to expand the volume 270 millimeter (10⅝ inch) diameter boreholes requires about twice the charge weight needed for 200 millimeter (8 inch) boreholes.

Should explosive weights be required that may cause wall damage and potential undercutting, a solution would be to split the charge in two, with one portion at the bottom and the other farther up the borehole, so an airbag could be placed at the desired locations and the upper charge of explosive using an airbag technology could be used to support the upper charge. An airbag could be placed at the desired location and the upper charge of explosive loaded on top.

The upper charge must be located next to a harder layer in the highwall.

Caution The upper and lower charges should be separated far enough to prevent damage to the wall by interacting shock waves.

Active Highwall Presplit Borehole Spacing

The tangential stress along a radial line from the boreholes must equal or exceed the force resisting cracking of the tensile crack to be successfully driven between presplit boreholes. The spacing must not exceed the relationship in equation 34.10.

$$S = \frac{4 \times P_b \times d}{\sigma_t + \sigma_n}$$ <!-- VERIFIED -->

Equation 34.10

Where:

• S = spacing between presplit boreholes (millimeters) (inches)
• d_e = borehole diameter (centimeters) (inches)
• P_b = the decoupled borehole pressure (kilograms/centimeter²) (pounds/inch²)
• σ_t = rock tensile strength (kilograms/centimeter²) (pounds/inch²)
• σ_n = rock tensile strength (kilograms/centimeter² (pounds/inch²)

For the borehole diameter in inches, and the borehole pressure and tensile strength in pounds/inch² the spacing will be determined in inches. When the diameter is given in millimeters, and the decoupled pressure and tensile strength are expressed in bars the spacing will be in millimeters.

The decoupled borehole pressure and therefore the weight of charge placed has an important bearing on the spacing between boreholes.

EXAMPLE 34.11

Calculate the borehole spacing using equation 34.10, having determined a suitable decoupled borehole pressure for the active highwall presplit design (1,020 bars). Testing has shown the tensile strength to be 81.6 bars (1,200 pounds/inch²) on average. The borehole diameter is 270 millimeters.

$$S = \frac{4 \times P_b \times d}{\sigma_t + \sigma_n}$$ <!-- VERIFIED -->

$$S = \frac{270 \times (1{,}020 - 81.6)}{81.6}$$ <!-- VERIFIED -->

$$S = 3{,}645$$ <!-- VERIFIED -->

The borehole spacing is 3,645 millimeters or 3.645 meters. Field experimentation is now required to optimize the design.

EXAMPLE 34.12

Calculate the borehole spacing using equation 34.10, having determined a suitable decoupled borehole pressure for the active highwall presplit design 15,000 pounds/inch². Testing has shown the tensile strength to be 1,200 pounds/inch² on average. The borehole diameter is 10.625 inches.

$$S = \frac{4 \times P_b \times d}{\sigma_t + \sigma_n}$$ <!-- VERIFIED -->

$$S = \frac{10.625 \times (15{,}000 - 1{,}200)}{1{,}200}$$ <!-- VERIFIED -->

$$S = 143$$ <!-- VERIFIED -->

The borehole spacing is 143 inches or 12 feet. Field experimentation is now required to optimize the design.

There are factors that limit the spacing between presplit boreholes even though the equation calculates a larger value. Joint or fracture spacing can be the controlling influence on presplit borehole spacing.

Caution Active Highwall Presplitting in Jointed Formations Generally, presplit borehole spacing for active highwall presplitting should not exceed 2 to 3 times the predominant joint spacing.

However, an active highwall presplitting widely spaced 3 meters (10 feet) or greater horizontal bedding planes are usually the predominant planes of weakness and this is usually not a problem.

In the case that equation 34.10 yields an ideal or larger spacing for the design dimensions selected. Field experimentation may demonstrate that these spacings do not lend to a high quality presplit. In that case the spacing will have to be reduced until an optimum result is determined by field testing.

The "4" in equation 34.10 is based on presplit experiments done in competent rock. Coal overburdens is a variable rock material and it is difficult to model exactly in many situations. Therefore, these computational techniques will produce reasonable initial results, which must then be optimized by field trials.

Caution Field testing is the most effective method of optimizing presplit borehole spacing for active highwall presplitting.

Implementation Of the Presplit

The horizontal bedding and pronounced plane of weakness at the coal/overburden interface means that a single charge fully coupled at the bottom of the presplit borehole is usually adequate. The two factors (1) angled boreholes, and (2) borehole spacing should be considered when implementing an active wall presplit blast.

Angled Boreholes

In practice, it is found that angled presplit boreholes work better than vertical holes. This approach accounts for any geological structure in the highwall that may undercut a vertical wall and allow blocks of material to slide out on the plane. Also, any undercutting that occurs near the bottom of the highwall is less likely to result in collapse of the wall above.

In Angled Highwall Presplitting (AHP), a common angle used is 70° (20° off vertical) leaving the new highwall at a 2.8:1 slope. Of course, angle presplitting requires the borehole drills to be capable of drilling on an angle.

Caution Each mine must assess the geologic structure information to determine the correct drilling angle.

When drilling angled presplit boreholes in variable topography, where the height of the highwall changes along the pit, adjustments must be made to obtain a straight highwall.

Presplit borehole locations must be the same at the toe.

In variable topography, this means that the collars of the presplit boreholes will not all be along the same line. The presplit line will not be straight but will vary according to where the boreholes pierce the surface of the topography.

Angled presplit boreholes are most effective with angled production boreholes. Then the burdens on the front row can be closely controlled from collar to toe leading to excellent relief of the toe and high castover percentages. Angle drilling of the production boreholes is quite common in modern coal mining.

Presplit Borehole Spacing

A method for calculating presplit borehole spacing is described in the Dragline AH-P Presplit Line section in this chapter. It is noted that geological structure has a dominating influence on the presplit borehole spacing in large diameter work, designs may need to be more conservative than the equation computes. Therefore, field experimentation and observations is always a component of optimizing hole spacings.

Common presplit borehole spacing varies from 3.05 meters (10 feet) to 5 meters (16 feet), depending on the strength and competency of the rock and the borehole diameter. Highly fractured, softer rocks require a closer spacing, while competent higher strength rocks (e.g., sandstones of compressive strength 82.74 megapascals to 172.4 megapascals (12,000 pounds/inch² to 25,000 pounds/inch²)), are satisfactorily presplit using a somewhat larger spacing. While these spacings are common they should be optimized for property by calculation and field trials.

Presplit Borehole Loading

A variety of methods for loading the large diameter presplit boreholes have been used. In the early days of active highwall presplitting bulk loaded detonating cord was often difficult. Another approach is to use a cross-linked packaged decoupled charge in specially designed tubes capped into the borehole. This has the advantage of allowing charges to be placed at various locations in the presplit borehole. However, this is time consuming and labor intensive.

By far, the most common technique now is to place a bulk loaded charge at the bottom of the borehole. This is easy to load and very labor efficient. Even in boreholes drilled at a 60° angle (30° off vertical) bulk loading usually works well.

If there is concern about coal damage some drill cuttings can be placed on the borehole bottom to provide standoff and cushion the coal from damage.

When the presplit charges in large, due to borehole diameter and depth, damage to the highwall in the vicinity of the charges may occur (See the size of the charge). The charge can be split into two and part of it placed farther up the borehole above a suitably placed air or foam bag. The total explosive weight remains the same.

When possible, wet presplit boreholes should be dewatered before loading. When the borehole cannot be dewatered, loading may need to be modified to limit the pressure transferred to the borehole wall. When the column of water is not too deep an air or foam bag could be placed at the top of the water and the explosive loaded on top.

Figures 34.25 and 34.26 shows splitting the explosive load to mitigate damage to the toe of the hole and to account for water in the borehole.

If the water column is long, one can still place a bag at the water surface. Also place some of the charge at the bottom of the borehole to ensure the lower portion of the hole is split. Some experimentation may be necessary to determine the amount of explosive weight to use in the bottom charge and that above the figures. Figures 34.25 and 34.26 shows splitting the explosive load to mitigate damage to the toe of the hole and to account for water in the borehole.

Presplit Borehole Stemming

In general, when column, the best results are obtained with unstemmed holes. The gases expand rapidly up the hole, provide momentary pressure to fracture between holes in tension and then vent and do no additional damage to the wall. Figure 34.27 is a photo of very successful presplitting at a coal mine with unstemmed 270 millimeter (10 5/8 inch) diameter boreholes.

In some cases, it is not possible to use unstemmed boreholes due to air overpressure effects on nearby neighbors. In this case the minimum stemming that controls the overpressure should be used. Stemming should be at the top of the borehole so that the detonation gases can expand fully up the presplit borehole, be reduced in pressure and provide a continuous presplit crack over the length of the borehole. A foam or air bag should be placed at the appropriate location and the stemming placed on top.

Relationship of the Production Back Row to the Presplit Line

A successful presplit will be ruined if the back row of production holes is placed too close to the presplit line. In other presplitting applications like open pit mines, quarries and surface construction it is usual to include a buffer row at the back of the blast that is lightly loaded and has an adjusted pattern.

In AHP it is not typical to use a buffer row because good fragmentation is needed for the dragline or a dozer to excavate the "keycut" material next to the new highwall. Therefore, it is important to carefully position the back row of holes to break the overburden but not damage the presplit.

Drilling both the presplit and production holes on the same angle is quite helpful in avoiding backbreak. Since front row burdens are very well controlled good relief and displacement toward the empty pit and not backward into wall is obtained.

When the presplit is angled and the production holes are vertically drilled it is necessary first to ensure a suitable offset at the toe between the two. Coming up the hole, the distance between the production hole and the presplit line continuously increases. At some point as slope overburden there will be too much burden between them and fragmentation and loosening will suffer. A row of shorter holes may be needed between the final full-depth production holes and the presplit to counteract the loss of fragmentation.

The amount of setback needed will depend on the overburden characteristics, the hole diameter and the explosive and is quite site specific. A good approach may be to make the initial offset fifty or sixty percent on normal production drilled burden. Then observe results and adjust accordingly.

Presplit Row Timing

AHP holes are typically shot before the production blastholes are drilled and shot. One advantage to this approach is that the presplit line can be observed after shooting which assists in the evaluation of presplit effectiveness. In dragline mines there may be some operational scheduling advantage as well. When the presplit line is initiated in advance of laying out and drilling the production boreholes, the presplit line should be surveyed in order that the distance between the presplit line and the last row of production holes is accurate with respect to design.

There should not be too much lag time between firing the presplit line and drilling and blasting the production boreholes. This will avoid uplifting the presplit crack with production hole debris due to flowing groundwater or runoff water which will degrade its effectiveness.

If the presplit line is shot with the adjoining cast the presplit holes must be detonated first. Enough time must be allowed for the presplit to effectively form. One hundred or more milliseconds are likely to be needed.

By contrast, if the presplit line leads by too much some disruption to back row holes of the production blast are possible. Therefore, a delay may need to be inserted into the line periodically to control the lag between the presplit firing and cast blast holes detonating.

It is advantageous to initiate as many presplit boreholes on a single delay as possible. However, in some cases ground vibration issues may limit the number of presplit boreholes that can be fired simultaneously. In these circumstances as many holes as possible should be detonated together and then a delay inserted.

EFFECTS OF CAST BLASTING ON DRAGLINE OPERATIONS

Dragline mining operations are varied in technique according to the depth of overburden, the number of mineable coal seams and the geometry of the dragline in use. Therefore, cast blasting techniques must be adapted to a variety of mining configurations. In addition, the results of cast blasting such as after blast profile affect how the dragline can be operated.

This section discusses aspects of how cast blasting and the dragline interact. It is not intended to present a thorough dissertation on dragline mining methods but rather to discuss the interaction between the dragline and the cast blast.

Single Seam Operation With The Dragline On The Highwall

Figure 34.28 shows a typical cross-section through a cast blast for a single seam operation and figure 34.29 shows a dragline mining cast overburden above a single seam after casting. Note the waste that the blast has displaced to its final location. Figure 34.30 indicates more than 60% of the material may be thrown clear of the coal under ideal conditions (McDonald, et al., 1962).

However, the amount of castover at any operation is site specific and dependent on geology, cast blast design and blast implementation. The percent total castover is related to the blast depth/width ratio as discussed earlier.

When the dragline operates from the highwall, there are two approaches. One method includes having the dragline mine in a conventional manner including removing the keycut and main cut portions of the blasted overburden. The second method has the dragline positioned to the outside of the cut and the keycut operation performed by dozers.

The after blast profile of cast blasts typically has the material positioned at a lower elevation than before the blast. In order for the dragline to position on the keycut it is necessary for the overburden to the side of the machine to be reduced in elevation to the dragline bench grade elevation. Otherwise the dragline cannot swing around to dump the blasted waste rock without striking the overburden to the side and damaging the machine. Figure 34.31 illustrates this approach for conventional highwall mining with casting included.

When the upper overburden is unconsolidated, tail room for the dragline can be produced using dozers. However, if the overburden is too hard to be moved by dozers it will be necessary to drill and blast to the depth of the dragline bench grade elevation before dozing the bench level.

The process of producing tail room for the dragline by blasting short boreholes to the side has generally proven difficult in practice. Therefore, while the conventional approach was with a cast blasting it has been largely given way to the second approach described below.

In this case, the dragline does not dig the keycut. Therefore, the use of cast blasting leads to the use of a bulldozer to cut the key due to the adjacent upper highwall. Therefore, the use of cast blasting leads to the use of a bulldozer or other equipment in the keycut becomes the best option.

Figure 34.32 shows a dragline digging in this fashion. The dozer that produces the keycut is seen in the photograph. Also, the lowering of the blasted overburden due to horizontal displacement can clearly be seen. It will be difficult to produce tail room on the side of the machine with this much drop, which is not uncommon. Therefore working the dragline on the crest and using a bulldozer or other equipment in the keycut becomes the best option.

Multiple Lift Operations

There is a wide variety of methods for mining multiple seams that employs draglines. When mining several coal seams the mining plan may call for the upper seams to be uncovered using shovel-truck methods. The waste above these seams is blasted in a conventional bench manner without casting.

Usually, only the overburden for first dragline mined seam is cast. It becomes difficult to cast additional overburden because material thrown into the pit from the first cast plus the dragline dug fill on top of the previously mined pit, making it difficult to achieve relief and throw more material into that excavation. A lower lift is subsequently blasted in a conventional manner, the dragline.

Similarly, very deep overburden above a single seam may need to be mined in two lifts. In this case the upper lift is cast. Figure 34.33 shows a dragline digging the upper lift that has been cast and leveled with dozers. Figure 34.34 shows leveling the cast material. The overburden below the tail room bench on the highwall will be drilled and blasted after the dragline encounters the upper lift.

When mining two seams with a thinner overburden and thick interburden it may be possible to cast the material above both seams. In this approach, the interburden is cast first. Then the overburden is cast down on the blasted interburden. Dozers then, level the blasted muck (See figure 34.34) to prepare the pad from which to operate the dragline.

Most of the overburden still must be handled by the dragline in this case. However, the dragline uncovers the upper coal seam by side cut and the lower seam by foot cut. Both seams are uncovered in one pass. This eliminates significant additional movement by the dragline that occurs if the seams are mined one above the other using two pit passes. In addition a percentage of the interburden is cast to find. Figure 34.35 shows a mine of this type. The dragline has uncovered the upper seam and is beginning the process of mining the interburden above the lower seam.

There are many methods by which draglines mine one or more seams. Even in the most complex methods cast blasting may well play an important role.

SUMMARY

Cast blasting is a viable method for increasing dragline pit production without needing to add costly new dragline capacity. Most of the increased cost in cast blasting is operating cost related to the use of higher powder factors. Capital cost outlays are usually modest.

A secondary benefit to cast blasting is better fragmentation and more loosening of the muck. The result is higher dragline production and reduced dragline consumable and maintenance cost.

Key considerations for designing cast blasts are listed in table 34.12.

Key Considerations For Designing Cast Blasts

Table 34.12 – Key considerations for designing cast blasts.

Caution Successful cast blasting requires a high level of quality control.

In reality, cast blasting is a specialized form of blasting for downstream benefit. Thus all the discussion of quality control in chapter 7 applies equally well to cast blasting.

Cast Blasting can make an important contribution to strip mines. In some specialized cases it may find a use in other types of open pit mines and quarries. Indeed there are large volume coal mines today that could not meet their production goals without the use of casting as a production technique.

REFERENCES

Chironis, N. P., 1984. Coal Age Operating Handbook of Coal Surface Mining and Reclamation, McGraw-Hill Inc. New York, New York.

Crosby, W.A. and A. Bauer. 1982. Wall control blasting in open pits, Mining engineering, p. 158, February. Society of Mining Engineers, New York, NY.

Livingston, Clifton. W. 1956. Theory of Fragmentation in Blasting, 6th Annual drilling and blasting symposium, University of Minnesota, October.

McDonald, K. L., W. K. Smith, and W. A. Crosby. 1962. Productivity improvements for dragline operations using controlled blasting in a single and multiple seam operation at Rietspruit, South Africa. Canadian Institute of Mining and Metallurgy, Annual Meeting, Quebec City, May.

Workman, J. L. 1995. The design, implementation, and optimization of casting blasts in strip mining. International Society of Explosives Engineers (ISEE) Proceedings of the 21st Annual Conference on Explosives and Blasting Technique, February 5 - 9, Nashville, TN. ISEE, Cleveland, OH.

Workman, Lyall and Peter Calder. 1991. A method for calculating the weight of charge to use in large holes pre-splitting for cast blasting operations. International Society of Explosives Engineers (ISEE) Research Proceedings of the 17th Annual Conference of Explosives and Blasting Technique, February 3 - 7, Las Vegas, NV. ISEE, Cleveland, OH.