Chapter 14: Blast Design Principles

The blast design defines the parameters required to break the rock with explosives and is based on the basic principles and parameters discussed in this chapter. Optimum parameters are difficult to determine initially as they depend on the site-specific characteristics of the rock at the blast site. All types of blasting apply the basic principles discussed in this chapter. The blasting chapters in Part III of the 18th Edition apply the principles in this chapter within the context of their discussions.

Efficient Blast Performance

Efficient blast performance requires effective utilization of explosive energy. Blast effectiveness is achieved when the five principal factors (1) energy distribution, (2) energy confinement, (3) energy level, (4) relief, and (5) explosive ratio are properly considered (See figure 14.1). In this chapter, these factors are not discussed in an order of importance.

Energy Distribution

The first principal performance factor is energy distribution. Uniform energy distribution in the rock mass produces uniform fragmentation. Each type of rock requires a minimum amount of explosive energy. Past blasting history and experience form the basis for choosing appropriate explosives.

The most significant blast design factor is the borehole itself. The borehole diameter limits the amount of explosive and consequently the total energy that can be loaded.

Caution

The borehole diameter limits the amount of energy that can be loaded.

Small diameter boreholes offer the advantage of smaller charges drilled on smaller patterns that disseminate the explosives energy more uniformly in the rock mass to maximize explosives energy distribution. This is an effective method to overcome the blasting limitations posed by highly fractured rock masses (See The Imperfect Rock Mass section later in this chapter).

Three other benefits of using small diameter boreholes are: (1) shorter stemming lengths to help to improve cap rock breakage, (2) reduced energy concentrations near the collar to help reduce the potential of flyrock, and (3) shorter subdrill lengths help reduce damage to floor below grade and beyond the advance surface in underground blasting.

In general, when the number of closely spaced smaller diameter boreholes increases the fragmentation size decreases. Conversely, when the number of larger diameter boreholes on larger pattern increases, the fragmentation size increases (Oriard, 2002).

A major disadvantage to small diameter boreholes is when the desired blasting production levels exceed the level that can be achieved by blasting with small diameter boreholes. For example, in a large quarry or surface coal mine that produces high volumes of broken rock to meet production goals larger boreholes may provide more benefit. (See chapter 7).

Energy Confinement

The second principal factor is energy confinement. Energy must be confined long enough for the detonation gasses to extend fractures and build pressure without prematurely venting. Konya (1985) summarizes the fragmentation and heaving process by the three-detonation phases: (1) Shock travels through the rock mass in all directions. It creates and extends cracks by compressional and tensile strain. Extension of these cracks is generally limited to a region of 20 to 30 charge diameters from the borehole; (2) The rising pressure of detonation gases forces the gasses into the existing cracks causing them to extend and form an expanded crack network; and (3) Gas pressure expands and causes the rock mass to bend, break and be displaced within a few milliseconds depending on the burden distance. A significant amount of fragmentation can be generated during this final phase. If the gas pressure is lost prematurely due to poor confinement, the fragmentation produced can be significantly reduced.

Explosive energy follows the path of least resistance. If the distance from the charge to the closest relief is shorter than the burden, confinement may be lost. Paths of least resistance are typically found in (1) low burden areas, (2) weak geological zones, or (3) weak stemming areas.

Energy Level

The third principal factor is energy level. Energy level refers to the total explosive energy applied to the rock. Each type of rock requires a minimum amount of explosive energy for fragmentation and rock mass movement. The total explosives energy loaded in a borehole depends on an explosive's energy per unit weight and its density. Considering an explosive's energy in addition to its density is a helpful way to (1) compare explosives of similar densities and (2) calculate the energy applied to a unit of rock (Postquick, 1999). In other words equal weights of two different explosives may contain different amounts of energy.

Commercial explosives manufacturers supply values of energy per unit weight and density for their products. Caution should be exercised when attempting to compare energies of different types of explosives (e.g. dynamite verses an emulsion) since different product types may have different energy basis calculations. For ammonium nitrate (AN) based products, different manufacturers may use different theoretical energy values for AN. Chapter 11 discusses accepted industry methods used to determine comparative (relative) energy values of similar types of explosives.

Relief

The fourth principal factor is relief. In this book "relief" describes the presence of a free surface (face) in the rock mass that borders a void space large enough for the blasted rock to occupy. Blast performance requires a free surface bordering a void of sufficient volume for the blasted rock to move and expand in volume. This surface is called the "free face". Key factors describing how relief and free face are given below.

Burden

Burden is the shortest distance from an explosive charge to its nearest free face (IME: SLP 12, 2010), making burden a critically important design factor.

Face

A face is a rock surface that marks the boundary of a rock mass of relatively uniform geotechnical properties. This can be a bench face, drilling surface, or rock structure or similar feature. Other common features (e.g. joints, faults, bedding planes, blasting damage, and geologic discontinuities) are imperfections that divide the rock mass, and affect the blast performance. Therefore, these features are not free faces because they do not border on relief. These rock imperfections affect blasting results in the two important ways by: (1) stopping the crack development of the detonation and (2) when they are weak they allow paths for energy to escape prematurely and reduce confinement.

Caution

Rock structures and mud seams have been referred to as free faces but neither provides relief for rock movement.

Free Face

Free faces are rock surfaces that border on relief. Free faces are usually the bench or drilling surfaces of the blast. Before blast initiation, free faces are illustrated in figures 14.2 and 14.3.

Relief

Relief is the presence of a void space of sufficient volume to accept the volume of broken rock from a blast. The void must be larger in volume than the volume of swelled rock produced by the blast. The location of relief is illustrated in figures 14.2 and 14.3. When more than one relief area exists as shown in figure 14.3, care must be taken to ensure rock movement is directed to the relief.

Subdrill

Subdrill is the additional borehole length drilled below grade level or advance distance to ensure rock breakage to the grade or advance surface (adapted from IME: SLP 12, 2010).

Swell

Swell is a measure of the bulk volume increase of a blasted rock mass to its original in situ volume. Blast swell is often in the range of 30%.

Figures 14.2 and 14.3 illustrate the free face(s) and relief prior to initiation. Free faces are also created during the detonation progression sequence where they provide ongoing relief during the blast. These "dynamic" free faces result from proper pre-blast relief, burden-spacing relationships, delay timing, and energy applied.

Dynamic free face formation can be visualized by drawing echelon lines on the delay diagram as shown in the series of illustrations in the Sequential Blasting section of this chapter.

Explosives Ratio

The fifth and final principal factor to consider is the explosives ratio. Explosive ratios are measures of how much explosives on a weight or total energy basis are applied to the rock in each borehole. Explosives ratios are commonly expressed as a powder or energy factors (See the Powder Factor vs. Energy Factor section of this chapter).

Yield

The overall explosive ratio is based on the borehole's yield. Yield is the amount of rock fragmented by the explosive over the borehole length to the grade or advance-final surface. The three dimensions used to calculate yield based on volume are (1) burden, (2) spacing, and (3) depth (See figure 14.4). Subdrill is often necessary to pull depth but is not considered to produce yield. Yield volume can be converted to weight by multiplying the volume by the rock's density. The following definitions are helpful when calculating the yield of a blast.

Spacing

Spacing is the distance between boreholes. In bench blasting, the distance is measured parallel to the free face and perpendicular to the burden (IME: SLP 12, 2010).

Bench Height

Bench height is the vertical distance from the top of a bench to the floor or to the top of the next lower bench (IME: SLP 12, 2010).

Depth of Advance

The depth of advance is the linear distance blasted in an underground blasting round such as found in tunnel, drift, raise, or shaft sinking operations (adapted from Rustan, 1998).

The metric and U.S. (sometimes called imperial) systems use similar terms to describe weight. Equivalent values are shown below:

U.S. - Metric Weight Equivalents

Yield volume is calculated using equation 14.1.

$$Y_{Vol.} = B \times S \times D$$ Equation 14.1

Where:

• $Y_{Vol.}$ = Yield volume (meters³) (yards³)
• $B$ = Drilled burden (meters) (feet)
• $S$ = Spacing (meters) (feet)
• $D$ = Depth of advance or bench height (meters) (feet)

EXAMPLE 14.1

Calculate the yield volume per borehole from a pattern with a burden of 5 meters, spacing of 6 meters and bench height of 20 meters using equation 14.1.

$Y_{Vol.(m^3)} = B \times S \times D$

$Y_{Vol.(m^3)} = 5 \times 6 \times 20$

$Y_{Vol.(m^3)} = 600$

The yield volume is 600 meters³.

In the U.S. system of measure, measurements are normally made in feet and volume is represented in yards³. The volume yield is converted to weight yield using Equation 14.2.

$$Y_{Wgt.} = Y_{Vol.} \times \rho_r$$ Equation 14.2

Where:

• $Y_{Wgt.}$ = Weight yield (metric tons) (tons) (See examples 14.2 and 14.4 below)
• $Y_{Vol.}$ = Volume yield (meters³) (yards³)
• $\rho_r$ = In situ rock density (kilograms/meter³) (pounds/yard³)

EXAMPLE 14.2

Convert the volume yield of 600 meters³ to metric tons for a rock with a density of 2,940 kilograms/meter³ using equation 14.2 and divide by 1,000 to convert kilograms to metric tons as follows:

$Y_{Wgt.} = \frac{Y_{Vol.(m^3)} \times \rho_r}{1,000}$

$Y_{Wgt.} = \frac{600 \times 2,940}{1,000}$

$Y_{Wgt.} = 1,764$

The weight yield is 1,764 metric tons.

EXAMPLE 14.3

Calculate the yield volume per borehole for a pattern with a burden of 16 feet, spacing of 20 feet and face height of 55 feet and divide by 27 to convert feet³ to yards³ using equation 14.1.

$Y_{Vol.} = B \times S \times D$

$Y_{Vol.} = \frac{16 \times 20 \times 55}{27}$

$Y_{Vol.} = 652$

The yield volume is 652 cubic yards.

EXAMPLE 14.4

Convert the yield volume of 652 yards³ to a weight volume for a rock with a density of 1,964 pounds/yard³ using equation 14.2 and dividing by 2,000 to convert pounds to U.S. tons.

$Y_{Wgt.} = \frac{Y_{Vol.(yd^3)} \times \rho_r}{2,000}$

$Y_{Wgt.} = \frac{652 \times 1,964}{2,000}$

$Y_{Wgt.} = 640.3$

The weight yield is 640.3 tons.

Powder Factor vs. Energy Factor

Powder factor (explosives loading factor) is the ratio of the weight of explosives used per unit of rock yield (adapted from IME: SLP 12, 2010).

In practice, powder factor can be reported in different ways as shown in table 14.1. The choice is usually a matter of local convention based on units used to evaluate expense and productivity. For example, projects based on volume removal such as in stripping and excavation projects use volume units (meters³ or yards³) where applications with performance based on weight of saleable product metric tons (tons) are generally used.

Table 14.1 - Typical powder factor units.

In contrast, energy factor means the total explosive theoretical energy applied to a unit of rock. It is normally represented in units of kilocalories per metric ton of rock. Energy factors can range from 100 kilocalories/metric ton to 450 kilocalories/metric ton (91 Kilocalories/U.S. ton to 408 kilocalories/U.S. ton). The energy factor must be high enough to overcome the rock's tensile strength. Borehole pressure must also reach sufficiently high levels for an extended period of time to provide proper expansion to move the broken rock.

Caution

Powder factor differs from energy factor. It accounts for only the weight of explosion used, whereas energy factor accounts for the total energy applied.

Types of Blasts

The two types of blasts (confined and face) differ only by the location of their relief. Efficient blast performance occurs when charges have sufficient energy to move rock into the desired relief. The other efficiency differs by the energy required to produce rock movement. Of necessity, confined blasts require higher explosives ratios than do face blasts. They must both separate the rock from the rock mass.

Confined Blast

Confined blasts are blasts where the relief is located above the borehole collar (See figure 14.2 in the Relief section of this chapter). Because of this, the blast must drive rock movement parallel to the borehole axis toward the collar using an adequate energy factor.

Cut

In underground blasting a cut is an arrangement of opening boreholes used in a round to quickly create relief to which the remainder of the round can break. In surface blasting a cut is defined to be the depth to which material is to be excavated to bring the surface to a predetermined grade (American Geological Institute, 2007).

Underground confined blasts (i.e. drift, tunnel, shaft rounds) are designed to create relief quickly. They have explosive ratios as the "cut" designed to eject rock quickly to create relief for the rest of the blasting round.

In the confined blasts described in chapter 33 (e.g. sinking and trenching blasts) the rock movement occurs within the confines of the blast site perimeter and is primarily due to rock swell rather than ejection. Chapter 33 discusses and illustrates typical delay pattern layouts for confined surface blasting applications. Since little of the rock moves outside this space, the explosives must be sufficient to separate the rock from the ground and break and continue moving or "fleezing" the previously blasted rock.

Independent of the application, confined blasts require relatively high powder factors and energy factors to move the rock to the relief.

Face Blast

Face blasts as discussed in this 18th Edition are those found in benching operations where opening boreholes are drilled parallel to the bench face that forms a boundary of the relief (See figure 14.3 in the Relief section of this chapter). Throughout the detonation sequence of a bench blast, new faces are continually created in each line of delays in the sequence as detonated and moves rock into its relief. It's easy to visualize how the successive echelons in the delay sequence illustrated in the Sequential Blasting section of this chapter create new faces as long as the previous echelon moves rock far enough to create an adequate relief.

Once the opening boreholes of any blast have moved rock into the relief, the blast can be thought of as a succession of face blasts. Therefore, the face blast is the most common type of blast. Initial and ongoing movement in face blasts is not as difficult to achieve as with confined blasts. When the bench face is determined to be true and vertical, the burden should vary little over the length of the face. Face blasts typically create movement to relief more easily than do confined blasts and do not require higher energy levels. Their energy must move the opening boreholes and successive free blasts of the dynamic free faces to provide sufficient relief for the final wall boreholes.

Pattern Layout for Sequential Blasting

Pattern layouts are arrangements of boreholes designed to provide the energy distribution for effective sequential blasting. Ultimately, pattern layouts must allow the driller to maneuver the drill and work safely. Pattern layout strategies differ between underground and surface blasts because of the borehole diameters typically used and techniques used to create initial rock movement.

Pattern layouts provide for an effective opening and "dynamic" face blast development throughout the delay sequence to the perimeter boundary.

Pattern geometry differs between surface and underground blasting. Surface pattern geometry is discussed and illustrated in the Bench Blasting Delay Pattern Layout section of this chapter to show how burden and spacing relationships are changed in the delay sequencing process when applied to the borehole geometry. Underground pattern and delay examples are illustrated and discussed in chapter 35.

Point Of Opening and Echelons

The borehole(s) at the "point of opening" must perform properly and not rob subsequent boreholes of their burden. The "point of opening" should be selected to ensure both (1) confinement and (2) proper burden. Subsequent boreholes and echelons are designed to break to the "point of opening" and create ongoing relief.

Surface Pattern Geometry

The uniform geometry of the square and equilateral triangle shapes results in uniform energy distribution within the blast site. This allows for effective echelon burden and spacing relationships to be developed in the delay sequence to produce the desired rock displacement (Workman, 2003). These geometric shapes are illustrated in figures 14.5 and 14.6 and are highlighted in the context of a layout pattern. Applications of these patterns are shown later in the Sequential Blasting section of this chapter. Unless angled boreholes are needed, face blast patterns typically use vertical parallel boreholes.

The square is a pattern drilled with equal burden and spacing dimensions where the angle between burden and spacing lines is 90°. Boreholes in successive rows are aligned directly behind those in the front or previous row.

In the equilateral triangle pattern boreholes are at equal distances from each other where the angle between lines forming the triangle is 60°. The drilled burden (measured perpendicular to the spacing) is no longer equal to the spacing but is 15% shorter.

The square pattern can be altered to a staggered-square without sacrificing the uniformity of energy distribution of the square. The staggered-square is laid out like the square but alternate rows are offset along the row at a distance equal to half the spacing. This variation creates different echelon angle possibilities.

If the square is altered to become a rectangle, the pattern loses its geometric uniformity because the spacing becomes longer than the burden. This geometric arrangement results in different echelon burden and spacing relationships than the square or staggered-square pattern. When the rectangle shape is considered, it should be carefully analyzed before blast design implementation.

Caution

Rectangular pattern layouts result in echelon burdens that may be too small (reducing confinement) and spacings too large to be effective.

Whatever geometry is used, the two overriding factors to consider are (1) the uniformity of energy distribution created and 2) the effective delay burdens and spacings created.

Underground Pattern Geometry

In underground blasting, boreholes are drilled both perpendicular to and angled to the drilling face, sometimes in the same pattern. Within an echelon or the "cut" they are parallel. They are parallel but have alignment problems optimum movement.

Angled boreholes are used to create better burden relationships with the drilling surface than boreholes perpendicular to the surface. Angled boreholes are used in V-cut, fan, and ring blast rounds.

Sequential Blasting

Proper delay sequencing is the key to muckpile formation and control of both vibration and flyrock. Since neither fragmentation crack extension nor movement occur instantaneously, timing is critical to maintaining the balance between required confinement and creation of ongoing relief.

Various rock types respond to the fragmentation process differently. Response time is also directly related to the degree of blast confinement. When blasts are more confined, they require more delay time in the sequence to assure movement.

Delay timing is designed to direct rock movement by creating echelons that direct displacement and create muckpiles. Proper timing also promotes borehole confinement and can minimize borehole-to-borehole interference.

Echelons

"Echelons" are groups of boreholes that form lines of contours of performance that create the ongoing "dynamic" free faces in the blast and direct muck movement. The distances between echelons in the blast pattern. Therefore, rock movement occurs in a perpendicular direction to the echelon line toward the free face as long as there is sufficient relief. The concept of echelon design is to create effective burden and spacing relationships for desired pavement direction. Spacing within an echelon could result in any one of the conditions listed in table 14.2.

Table 14.2 - Echelon spacing performance factors.

As a general rule, the maximum spacing between boreholes within an echelon should not exceed twice the burden, since the burden represents the limit of fragmentation cracking. In addition, rock movement velocity decreases as the burden increases making burden control critical to the ongoing creation of relief. Burden movement velocity is dependent on the rock type (Chiappetta, 1991).

Delay Timing Strategies

The strategy of delay timing is to provide appropriate time delay within and between echelons for rock fragmentation and movement. This emphasizes the need to design a proper burden and spacing relationship with echelons. Timing between echelons is important to the ongoing creation of relief. The delay time and burden at the blasting boundary needs to be sufficient to establish and protect the final wall. Up to a point, time intervals between echelons are greater in confined blasts than in face blasts to create the ongoing relief required for total movement and the preservation of the blasting boundary.

The following bench blasting delay pattern layouts illustrate how the square and equilateral shapes can be used in bench blasting applications to produce burden and spacing relationships for direct movement. Since underground patterns are drilled into vertical or overhead boundaries they do not have the same degree of latitude to be altered as in the bench face blasting layouts. Underground patterns and delay are illustrated and discussed in chapter 35.

Bench Blasting Delay Pattern Layout

Figures 14.8 through 14.16 illustrate various echelon burden and spacing relationships using the square and equilateral triangle drill pattern geometries in various bench blast configurations. Since rock response depends on many factors these diagrams illustrate only echelon designs to gain burden and spacing advantages. The delay timing is left to the blaster-in-charge to determine based on the specific rock type and condition at the blast site, the blasting goals, and the resulting echelon burden. Guidelines to create the echelon angles shown in figures 14.8 through 14.16 (POI indicates point of initiation) are outlined in table 14.3 for delay pattern layout for this series of figures produce echelon angles to the free face to direct movement as illustrated.

The guidelines in table 14.3 can be applied to any regular pattern layout. However, effectiveness is limited by the echelon angle formed with the face. As echelon angle increases, forward muck movement decreases. Echelon angles should be calculated to ensure movement is properly directed.

Table 14.3 - Guidelines for echelon design for bench blasting (See Figures 14.8 through 14.16).

The Imperfect Rock Mass

Geologic structures, other fracture planes, voids and any other imperfections in the rock mass limit blast performance in three ways when they (1) inhibit full fragmentation crack network formation, (2) enhance unwanted rock movement such as flyrock, and (3) interrupt the uniform energy distribution provided by the pattern. Limiting factors resulting from these rock features are summarized in table 14.4. These rock features are discussed and/or illustrated in chapter 8.

Table 14.4 - Rock features and their limiting factors.

Pattern Row Orientation Strategies

The blast designer should orient patterns to minimize the effects caused by geologic structures. Figure 14.17 illustrates the best face orientation in bench blasting relative to the strike of bedding planes. The face is perpendicular to the strike.

Figures 14.18 and 14.19 illustrate row orientation parallel to the strike of dipping bedding planes.

Figure 14.18 illustrates an orientation that favors backbreak. When the bedding planes are weak, beds often slide into the pit and cause cutoffs of unconnected downlines. Figure 14.19 illustrates better conditions for backbreak control, but grade may be difficult to pull if energy escapes up dip at the borehole bottom.

Overriding Geologic Factors

Structure spacing is typically a good predictor of blast fragment size if it causes nonuniform energy distribution. Poor explosives energy distribution occurs when pattern dimensions are wider than structure spacing.

Caution

Poor energy distribution is the result of a failure to get energy into in situ blocks.

Unless patterns are designed to better distribute explosives energy evenly within the blocks, fragmentation will be limited to that obtained in the mass rock movement of the blast. In addition to structures, blast damage and voids in the rock often become overriding performance factors. If not addressed in the design to provide proper borehole burden and confinement, blast results may be affected.

Structures should be identified and mapped. Figure 14.20 illustrates a pattern too large for the joint spacing. This distribution often occurs when larger patterns are used with larger boreholes. Figure 14.21 illustrates better energy distribution when using smaller boreholes on smaller patterns.

The presence of structures can sometimes be advantageous. When structures occur at grade or the end of an advance and are weak, they can reduce the need for subdrilling. Mine development plans or project development plans should locate bench heights or depths of advance to end at these locations whenever possible. Structures in these locations may also help control overbreak, causing one definition in these areas.

Column Loads

Boreholes are normally loaded as a single continuous charge of uniform energy level. Some applications require the column to be split to preserve confinement through weak geologic conditions or to control ground vibration.

Continuous vs. Split

Continuous column load energy level can be varied to meet variable rock conditions by both conventional and bulk loading methods. Bulk loading with gassed emulsion products provides varying energy from the borehole bottom to the stemming (See chapter 11). Special precautions should be heeded when using chemically gassed sensitized products, and the manufacturer is the best source of advice. When continuous columns are divided into discrete energy zones using different products, the blaster-in-charge must ensure each product will adequately prime the next product loaded.

Split columns physically divide the column into discrete charges using an inert material ("decks") to separate the charges. This is usually the stemming material. The uncharged charges in a split column must be individually primed. In split column the charges can be either initiated at the same time or delayed. When delay times are required between adjacent charges, the stemming interval should be sufficiently long to prevent the first charge from damaging, uncoupling, or influencing the second charge. Wet boreholes need longer stemming intervals.

Caution

Wet boreholes require stemming of sufficient length and quality to preserve explosion confinement and prevent interference between adjacent charges.

Split columns and decks are sometimes used in deep and highly confined surface applications to create top to bottom delay sequencing to enhance relief in a highly confined bottom area.

Coupled vs. Decoupled

Optimum fragmentation depends on efficient energy transfer to the borehole wall.

Maximum transfer occurs in fully coupled charges. Figure 14.22 illustrates borehole coupling. The success of some applications such as presplitting is based on decoupled charges that limit energy transfer. In an application like presplitting, maximum fragmentation is not a blasting goal.

Coupling

Coupling is the degree to which an explosive fills the cross section of a borehole (IME: SLP 12, 2010). Bulk-loaded explosives are completely coupled, while untamped rigid packaged explosives are decoupled. Untamped non-rigid packaged explosives slc, however, slump somewhat under the column weight, which increases their degree of coupling.

When loading cartridged or packaged explosives, decoupling occurs to some degree. Cartridges in plastic wraps will tend to slump and couple better. Cartridges in rigid shells remain uniformly decoupled since they do not slump. It is possible to calculate the amount of the borehole filled with cartridged or bagged explosives calculating the "coupling ratio", discussed later in the Blast Design Parameters section of this chapter.

Coupling is also a term used to describe the continuity of the explosive charge within the borehole.

Caution

Gaps within the column load negatively affect and possibly prevent the detonation reaction from progressing from the primer to the ends of the charge. Gaps may sometimes occur during the loading process.

Cratering Principles

Cratering is the fragmentation process of a single buried charge is discussed in chapter 9. An understanding of cratering performance aids blast designers to properly consider the potentially varying results that may occur borehole column, two- and multi-ball issues. Despite the fact that cratering theory is based on the performance of a point spherical charge, pointed and cylindrical charges may perform like point charges. This occurs in the oddveall and collar portions of the borehole. It is helpful for the blast designer to visualize the cratering process as illustrated in figure 14.24 when designing for these areas. In addition, the cratering process is one that affects the effective burden to spacing relationship.

Charge Shape

While research and practice offer various findings, for practical purposes, the cylinder charge shape performs almost like that of a sphere when its length is less than, or equal to, 8 times the borehole diameter as illustrated in figure 14.23.

Where:

• $d_e$ = charge diameter (millimeters) (inches)
• $l_c$ = cylindrical charge length measured in same units as $d_e$.

EXAMPLE 14.5

Determine if a 375 millimeter charge 3 meters long will perform like a spherical charge of the same weight using the relationship in figure 14.23.

Step 1 Convert charge diameter to meters (same units as length) by dividing by 1,000 as follows:

$D_{e(m)} = \frac{d_{(mm)}}{1,000}$

$D_{e(m)} = \frac{375}{1,000}$

$D_{e(m)} = 0.375$

The charge diameter is 0.375 meters.

Step 2 Calculate the length to diameter ratio using the relationship in figure 14.23.

$R = \frac{l_c}{d_e}$

$R = \frac{3}{0.375}$

$R = 8$

The ratio is 8. Since this ratio satisfies the condition (≤8) in figure 14.23, the charge's performance will approximate that of a spherical charge.

EXAMPLE 14.6

Determine if a 6.5 inch diameter charge 8 feet long will performs like a spherical charge of the same weight using the relationship in figure 14.23.

Step 1 Convert charge length to inches by multiplying by 12 as follows:

$l_{c(in.)} = l_{c(ft.)} \times 12$

$l_{c(in.)} = 8 \times 12$

$l_{c(in.)} = 96$

The charge length is 96 inches.

Step 2 Calculate the length to diameter ratio using the relationship in figure 14.23.

$R = \frac{l_c}{d_e}$

$R = \frac{96}{6.5}$

$R = 15$

The ratio is 15. Since this ratio does not satisfy the condition (≤8) in figure 14.23, the charge's performance will be that of a cylinder.

Effective applications of cratering principles at a specific blast site require trials to determine the rock cratering performance characteristics. Site-specific studies have been used to verify common burden, burden and spacing relationships, stemming lengths and toe and collar breakage. Such studies produce data on crater that can be considerable and make interpretation difficult. Therefore, a better approach is normally to refine burdens and spacings based on the previous blast design results (Workman, 2003).

Terminology

Craters are described by the terms in figure 14.24. The Fragmentation and Heave Process chapter discusses and illustrates the cratering process. In vertical upward cratering movement, rock fully ejected from the true crater falls back and creates the apparent crater. In the inverted vertical cratering application of the vertical crater retreat blasting (See chapter 35) the fall back into the true crater does not occur.

The five important terms used to describe the final crater perimeters are the (1) apparent crater, (2) lip, (3) radius, (4) depth of burial (DoB), and (5) true crater.

Apparent Crater

The apparent crater includes the true crater plus all the fallback material. The curved apparent crater profile is sometimes best approximated by a hyperbola.

Radius

The radius is the distance from the charge ($R_{cr}$) where the true crater intersects the surface.

Lip

The lip is formed by broken material heaved and ejected at and outside the circumference of the true crater. The lip resembles a berm.

Depth of Burial

The depth of burial (DoB) is the depth from the original surface to the center of the charge.

True Crater

The true crater represents the profile made where all the material surrounding the charge has been fragmented. It is the crater that would be observed if there was no fallback into the excavation, i.e. if all the material was removed after detonation. The true crater is described by its depth, radius and volume.

The crater formation is depicted in the sequence of illustrations in figure 14.25. This is a dynamic process, since the rock responds to the detonation over time (Workman, 2003).

Scaling Factors

Cratering studies result in two sets of site-specific data that can be correlated including the (1) type and energy of the explosive and the (2) rock mass geotechnical properties at optimum conditions and can produce accurate cratering parameters for blast designing.

The most common method of predicting crater dimensions and volume is by empirical scaling. Methods are also available that apply the laws of conservation of mass, momentum and energy. However, the computations are very complex and are better suited to researchers and computer modeling. The two factors most responsible for crater performance are (1) the charge shape and (2) its depth of burial.

Charge Shape

Spherical charge performance is related to the inverse of the "cube" root of the optimum charge weight at the optimum DoB. By contrast, cylindrical charge performance is related to the inverse of the "square" root of the optimum charge weight at the optimum DoB. This relationship helps explain why the cylinder shape charge produces a better fragmentation effect.

Depth of Burial

Depth of burial (DoB) is the burden for a cratering charge and significantly affects crater performance.

Actual Depth of Burial

The actual DoB is the distance the charge is buried below the ground surface (free face). Therefore, the actual DoB is the burden.

Critical Depth of Burial

The critical DoB is the depth beyond which no crater is formed and there is no ejection or rock heave at the surface. The only breakage phenomena at this depth are crushing around the charge and some radial cracking in the vicinity of the charge.

Optimum Depth of Burial

At optimum DoB, the largest crater is produced (See figure 14.24). When the DoB is shallow, high sput velocities are generated. This results in flyrock and smaller crater dimensions. At shallow depths the high-pressure gases vent through the surface so rapidly that there is little time for them to exert pressure on the ground.

As the charge is buried progressively deeper than optimum in the same material, the crater dimensions decrease. Spalling related to the tensile wave is confined to a layer near the face. Gas acceleration effects on the rock are countered by the weight of the material.

Scaled Depth of Burial

Empirical scaling is the most common method of producing crater dimensions. Since crater performance directly depends on the nature of the rock material and the explosive type used, the optimum DoB is most effectively found from site-specific cratering studies. Once site-specific data is compiled, cratering predictions are more reliable and appropriate scaling factors can be applied.

Field Cratering Studies

Cratering predictions will be more reliable and appropriate scaling factors can be applied when site-specific data is compiled and analyzed. The value of site-specific crater studies is that they can help verify the seven important pattern and performance factors listed in table 14.4.

Table 14.4 - Design and site factors verified by cratering studies.

Blast Design Parameters

Blast designs are based on basic parameters discussed in this chapter. Blast performance is the result of properly designed parameters and the factors discussed at the beginning of this chapter. Today, technology tools and methods are available to determine and measure these parameters more exactly. When fully considered, blasting results will better match those that are expected. These parameters are interrelated. When one is changed others will need to be changed. Based on this interrelationship there is a practical limit for any blast advances, and the limit is based on the borehole diameter.

Borehole Diameter

Borehole diameter controls the amount of explosive energy that can be loaded. The diameter is then the basis for the burden (See equation 14.3) and ultimately the pattern layout dimensions. Accuracy of the diameter is a function of drilling quality. Slight variations in diameter caused by wear (smaller diameter) and geologic structure (larger diameter) can cause significant variations in the borehole diameter and consequently the amount of explosive loaded. Therefore, borehole diameter directly affects fragmentation results.

Borehole Length

The desired blast depth (bench height or advance depth) along with required subdrilling determines the borehole length. In many applications extra subdrill length must be added to pull the grade or advance distance. The deck or stacking quality dictates the amount of subdrill required in relation to the burden.

Burden

Burden is the distance of the explosive charge to the nearest free face of the relief (adapted from IME: SLP 12, 2010). The optimum burden is the distance of radial cracking produced by the detonation to produce the desired fragmentation or movement. Therefore, burden is both a pattern layout dimension and an explosive performance factor. The echelon burden should not exceed the performance capabilities of the explosive.

If the burden is less than the radial cracking radius produced by the detonation, energy venting and excessive movement from the face can occur. If the burden is significantly greater than the cracking radius, insufficient fragmentation and movement result. Therefore the borehole diameter controls the burden. After the borehole diameter is chosen, the next parameter to determine is the burden. (Oriard, 2002).

If the burden distance is excessive in a box blast, the path of least resistance could be directed toward the borehole collar.

Roy and Singh (1998) and Pugliese (1982) each summarizing various empirical equations to calculate the approximate burden for bench blasting. Such equations are typically based on many site variables and factors. When using such equations the reader is cautioned to consider all stated applicable factors and stated limitations that affect the equation's use. The empirical equation 14.3 was given by Konya (1998) and is presented here.

$$B = \left(\frac{2 \times SG_e}{SG_r} + 1.5\right) \times D_e$$ Equation 14.3

Where:

• $B$ = Burden (feet)
• $D_e$ = Diameter of explosive (inches)
• $SG_e$ = Specific gravity (explosive)
• $SG_r$ = Specific gravity (rock)

EXAMPLE 14.7

Calculate a design burden for a quarry bench blast in dolomite (specific gravity = 2.5) using a 165 millimeter diameter borehole and an emulsion (specific gravity=1.2).

Step 1 Convert the borehole diameter to inches. Multiply by 0.0394.

$d_{(in.)} = d_{(mm.)} \times 0.0394$

$d_{in.} = 165 \times 0.0394$

$d_{in.} = 6.5$

The diameter is 6.5 inches.

Step 2 Calculate the burden using equation 14.3.

$B = \left(\frac{2 \times SG_e}{SG_r} + 1.5\right) \times D_e$

$B = \left(\frac{2 \times 1.2}{2.5} + 1.5\right) \times 6.5$

$B = \left(\frac{2.4}{2.5} + 1.5\right) \times 6.5$

$B = 2.46 \times 6.5$

$B = 16$

The burden is 16 feet. When converted to metric it is 4.9 meters.

Spacing

Spacing is the distance between boreholes in a row always measured perpendicular to the burden (IME: SLP 12, 2010). In the drill pattern geometries illustrated in the Bench Blast Delay Pattern Layout section of this chapter, the delayed spacing is greater than or equal to the drilled burden for most blasting operations.

Caution

Careful consideration should be given to designs where the spacing is shorter than the burden.

When the spacing is greater than the burden, it is easier to create a rock movement in the burden direction.

Subdrill

Subdrill is the added length distance that the borehole is drilled past the desired grade or advance plane. Where a geologic structure occurs at the grade or advance plane location, subdrill may not be necessary if the structure is weak. In other cases, some subdrill length is required to ensure maximum yield is obtained.

If subdrilling is required, it is important to limit it to only the minimum length required. Excessive subdrilling causes floor damage in benching operations and face damage in underground projects. It may also cause higher ground vibrations from this confined location underground. Performance in the subdrill is based on the effectiveness of the cratering performance. Optimum subdrill lengths can be verified by a site-specific cratering study. In the absence of such a study, a good industry practice is to calculate the subdrill length by multiplying the burden by 0.2 to 0.5. The conservative starting point of 0.2 can be increased on the basis of blast evaluation and a review of floor lips.

Ideally, the subgrade center of each borehole illustrated in figures 14.26 through 14.28 intersect at grade. The blaster-in-charge should expect that the depth of loose rock to drill through on the next bench will be approximately equal to the subgrade drilled.

Stemming

Stemming is the inert material used to confine energy within the borehole at the top of the explosive charge. Insufficient stemming lengths allow energy to prematurely vent at the collar and can contribute to flyrock. Excessive stemming lengths reduce energy distribution in the area near the collar. Empirical cratering data can be used to verify optimum stemming length. In the absence of cratering data, a common industry practice is to calculate the stemming height by multiplying the burden by 0.7. Stemming lengths for future blasts should be properly modified after the post blast evaluation.

Stemming lengths in face blasts can range between 20 and 30 charge diameters depending on the conditions of the site. Optimum stemming lengths are very site-specific and are directly related to the factors in table 14.5.

Table 14.5 - Stemming length performance factors.

Caution

Explosive charges should never be placed in weak material or fill material above the solid rock.

It is recommended that drillers record the borehole conditions on a borehole-by-borehole basis and provide this information to the blaster-in-charge so boreholes can be properly stemmed (See chapters 18 and 19).

Clean crushed rock chips sized to approximately ½ the borehole diameter should be used for stemming material when good fragmentation of the upper portion of the bench is desired. This allows the stemming length to be maintained without losing energy confinement. Stemming length should be cautiously adjusted after the blast evaluation. Wet drill cuttings provide less locking capacity.

Caution

When clean crushed chips are not available and drill cuttings must be used, stemming lengths should be longer as cuttings do not lock in the borehole as well.

Stemming plugs like that illustrated in figure 14.29 can be used to improve borehole confinement by helping to lock loaded stemming into position.

Loading Density

Loading density is a measure of the total weight of explosives in a unit length of the borehole and a measure of the energy loaded. Equal weights of different density explosives yield different loading densities. Loading density and the relative explosives strength calculations discussed in chapter 11 form the basis for relative scaling of blast designs based on a change in explosive type. Approximate loading density can be found by calculation using either equation 14.4 (metric) or 14.5 (U.S.).

Caution

Loading density chart values, if available, are not as accurate as those obtained by calculation.

Loading density directly depends on the three factors: (1) borehole diameter, (2) product density, and (3) degree of coupling. Since the diameter is squared in equations 14.4 (metric) and 14.5 (U.S.), it follows that a slight variation in borehole diameter produces a significant variation in the loaded density. Be aware that borehole diameter varies as a result of both (1) drill bit wear; and (2) rock quality (soft or hard), and its in situ fractured characteristics.

$$\rho_{ld} = 0.000785 \times d_e^2 \times \rho_e$$ Equation 14.4

Where:

• $\rho_{ld}$ = Loading density (kilograms/meter)
• $d_e$ = Diameter of the explosive (millimeters)
• $\rho_e$ = Explosive density (grams/centimeter³)

EXAMPLE 14.5

Calculate the loading density of a bulk explosive with a 1.23 grams/centimeter³ density in a 170 millimeter diameter borehole using equation 14.4.

$\rho_{ld} = 0.000785 \times d_e^2 \times \rho_e$

$\rho_{ld} = 0.000785 \times 170^2 \times 1.23$

$\rho_{ld} = 27.90$

The loading density is 27.9 kilograms/meter.

In the U.S. system the loading density is calculated using equation 14.5.

$$\rho_{ld} = 0.3405 \times d_e^2 \times \rho_e$$ Equation 14.5

Where:

• $\rho_{ld}$ = Loading density (pounds/foot)
• $d_e$ = Diameter of the explosive (inches)
• $\rho_e$ = Explosive density (grams/centimeter³)

EXAMPLE 14.6

Calculate the loading density of a bulk explosive with a 1.23 grams/centimeter³ density in a 6.75 inch diameter borehole using equation 14.5.

$\rho_{ld} = 0.3405 \times d_e^2 \times \rho_e$

$\rho_{ld} = 0.3405 \times 6.75^2 \times 1.23$

$\rho_{ld} = 19.1$

The loading density is 19.1 pounds/foot.

Caution

Borehole diameter variations may result in inventory discrepancies when blast log weights are based on the loaded density calculation. When bulk loading emulsion, attention must be paid to bit wear, geologic structure, and drilling technique as these factors can affect borehole diameter and cause significant accounting errors.

The issues resulting from loading density values are summarized in table 14.6 and should be considered.

Table 14.6 - Loading density issues.

Coupling Ratio

Coupling ratio is the ratio of the cross-sectional area of an explosive charge to the cross-sectional area of the borehole (Rustan, 1998). It is commonly used as an approximate indicator of how well energy will be transferred to the borehole wall. Full contact provides the most efficient energy transfer. There are applications such as presplitting or smooth wall blasting where full coupling is not a desirable parameter. As well, there are specialty blasting considerations that utilize variations of coupling or decoupling techniques in order to achieve optimum results.

$$K_c = \left(\frac{d_e}{d_b}\right)^2$$ Equation 14.6

Where:

• $K_c$ = Coupling ratio (dimensionless)
• $d_e$ = Diameter of explosion (millimeters) (inches)
• $d_b$ = Diameter of borehole, consistent units with $d_e$ (millimeters) (inches)

EXAMPLE 14.7

Calculate the coupling ratio of a 88 millimeter diameter cartridge explosive loaded in a 102 millimeter diameter borehole using equation 14.6.

$K_c = \left(\frac{d_e}{d_b}\right)^2$

$K_c = \left(\frac{88}{102}\right)^2$

$K_c = 0.744$

The coupling ratio is 0.744. This means the borehole is approximately 74.4% filled.

EXAMPLE 14.8

Calculate the coupling ratio of a 4.0 inch diameter cartridge explosive loaded in a 5.0 inch diameter borehole using equation 14.6.

$K_c = \left(\frac{d_e}{d_b}\right)^2$

$K_c = \left(\frac{4}{5}\right)^2$

$K_c = 0.64$

The coupling ratio is 0.64. This means the borehole is approximately 64% filled.

Decking

Decking is a method of creating unloaded zones within the borehole to enhance explosive performance or limit the charge weight initiated at any given time. Enhancement may be improved by either isolating the energy from a weak geologic zone or by creating an unloaded zone to cushion the final wall. When used to separate charges for independent performance, sufficient inert length must be provided to prevent charge-to-charge interference that can result in explosive malfunction. These separations are also known as decks. Specific uses of decks are summarized in table 14.7. Various decking use techniques are illustrated in figure 14.30.

Table 14.7 - Summary of decking uses.

If the deck separates two charges so they detonate at different times, it is common that the deck length be approximately 14 charge diameters. This separation distance helps avoid malfunction of the later firing charge.

Air decking is used to reduce the charge weight in the borehole while maintaining the overall energy distribution within the bench. Air decking also allows the borehole pressure to be a leverage mechanism in final wall blasting situations. One advantage of axial air decks is that the explosive column remains continuous, so additional in-hole delays and boosters are not required.

Stiffness Ratio

Hustrulid (1985) recorded rock flexure along the borehole axis and concluded that face flexure contributed significantly to rock breakage. Flexure was greatest toward the middle of the borehole length, and flexure was determined to depend on the borehole's stiffness ratio (height-to-burden ratio) as denoted in equation 14.7. Therefore, the stiffness ratio is an indicator of the relative ease at which a bench face breaks and moves along its height profile. In general, patterns with low stiffness ratios don't displace rock as far, whereas those with a higher ratio displace more. Movement distance is at the middle of the face and least near the floor or grade line. Rock stiffness is dependent on its strength and elastic properties.

Table 14.8 lists probable blasting outcomes for various stiffness ratio values. Based on rock type the results may vary.

Table 14.8 - Probable blasting outcomes for various stiffness ratio values. (Courtesy: Lilly Explosives Company, 1992)

$$R = \frac{H}{B}$$ Equation 14.7

Where:

• $H$ = Face height (meters) (feet)
• $B$ = Burden (meters) (feet), units consistent with H

Example 14.9

Calculate the stiffness ratio for a face 10 meters high and a 2 meter burden using equation 14.7.

$R = \frac{H}{B}$

$R = \frac{10}{2}$

$R = 5$

The stiffness ratio is 5. This indicates there may be poor toe movement or fragmentation.

Example 14.10

Calculate the stiffness ratio for a face 35 feet high and a 6.5 foot burden using equation 14.7.

$R = \frac{H}{B}$

$R = \frac{35}{6.5}$

$R = 5.4$

The stiffness ratio is 5.4. This indicates there may be poor toe movement or fragmentation.

Cap Rock

Cap rock refers to a solid layer of rock within the stemming area. Since explosive column loading is generally controlled in the stemming zone for flyrock control and energy confinement reasons, cap rock is at risk of not being broken by the main column charge. One technique to provide cap rock breakage is to load small decoupled high brisance charges placed inside the cap rock layer in regular patterns or in extra smaller boreholes ("satellite boreholes") drilled between the pattern as the cap rock as illustrated in figure 14.32. Cap rock charges are designed to crack only the cap rock. Since forward movement in the stemming zone is relatively low when compared to the face, broken cap rock chunks are often found on top of the muckpile.

Stand Off

Stand off refers to a drilling and loading practice to locate the bottom of the charge at a distance from the toe to prevent ore damage or dilution. Stand off is commonly used to prevent coal damage when blasting overburden or interburden above coal seams.

The length of stand off depends upon the strength of rock being blasted and strength of the ore potentially affected. The borehole diameter controls the energy in the bottom of the borehole. In coal mining, the standoff can range between 1 meter and 10 meters (3 feet to 33 feet).

Angled Boreholes

The two primary reasons to use angled boreholes are to either (1) enhance rock movement or (2) result in final walls at specified angles. For the two reasons stated, maintaining angle accuracy is critical to ensure the performance they are to provide. This section discusses the methods and calculations used to determine the measurements necessary to determine borehole parameters necessary for borehole layout and measurement.

Underground blasting (See chapter 35) may use angled boreholes in cuts, and in ring blasting. In the V-cuts, they are used to create initial relief and provide for the ejection of rock necessary for ongoing relief. In ring blasting they provide for energy distribution in a large rock mass when drilling from a single point location in a caving operation (See chapter 35). Sometimes boreholes are angled slightly to provide for waste damage or perimeter relief in confined blasting environments. In these cases, the driller usually makes the angle adjustment for the slight angle involved.

Surface and final wall blasts (including cast) (See chapters 33 and 34) often incorporate angled boreholes in the geometry. Slant blasting begins to create an a final wall at a specified slope. Examples are: (a) for the specified final slope angles found in some road cuts or excavations. Bench blasts use angled boreholes to overcome excessive face burden conditions (arching too).

Cast blasts utilize angled boreholes to create sloping highwalls for (1) final face stability and (2) productive benchless for cast ore blasting.

If using angled boreholes, both the (1) angle of the borehole and (2) its length must be known. Where face profiles are made they may indicate the position of the bottom for design determination. In bench blasting the best method drill and length angle can be chosen to achieve a practical borehole as shown in fig. 14.33.

Based on the desired position of the borehole bottom, an angled borehole must be collared at a horizontal offset from the bottom as shown in figure 14.33. The series of equations 14.8 through 14.10 are used to calculate the unknown quantity when only two are known. Equation variables are listed following equation 14.8 only. A face profile as discussed in chapter 32 is indispensable when determining these measurements. The offset is calculated from the borehole angle using equation 14.8.

$$O = D \times \tan(A)$$ Equation 14.8

Where:

• $O$ = Collar offset (units same as those for $H_b$)
• $H_b$ = Bench height (any units)
• $D$ = Borehole depth (same units as $H_b$)
• $l$ = Borehole length ($H_b$ + subdrill, same units as D)
• $A$ = Drilling angle (degrees)

When the borehole length must be calculated, this equation can be written as in equation 14.9.

$$l = \frac{O}{\sin A}$$ Equation 14.9

The angle can be calculated from the offset by using equation 14.10.

$$A = \tan^{-1}\left(\frac{O}{D}\right)$$ Equation 14.10

EXAMPLE 14.11

Calculate the offset required for a 10° angled borehole 16 meters long, using equation 14.9.

$O = l \times \sin A$

$O = 16 \times \sin 10$

$O = 16 \times 0.174$

$O = 2.8$

The offset from the crest is 2.8 meters.

EXAMPLE 14.12

Calculate the offset required for a 10° angled borehole 53 feet long using equation 14.9.

$O = l \times \sin A$

$O = 53 \times \sin 10$

$O = 53 \times 0.174$

$O = 9.2$

The offset from the crest is 9.2 feet.

EXAMPLE 14.13

Calculate the borehole angle and length where the bench height ($H_b$) is 14 meters with a required subdrilling of 1.0 meters and the horizontal offset off the crest to toe by survey is 3.6 meters.

Step 1 Calculate the depth of the borehole bottom from the collar as follows:

$D = H_b + l_{subdrill}$

$D = 14 + 1$

$D = 15$

The depth of the borehole bottom from the collar is 15 meters.

Step 2 First calculate the required angle using equation 14.10.

$A = \tan^{-1}\left(\frac{O}{D}\right)$

$A = \tan^{-1}\left(\frac{3.6}{15}\right)$

$A = \tan^{-1}(0.24)$

$A = 13.5$

The drill angle is 13.5°.

Step 3 Calculate the borehole length using equation 14.9.

$l = \frac{O}{\sin A}$

$l = \frac{3.6}{\sin(13.5)}$

$l = \frac{3.6}{0.233}$

$l = 15.4$

The borehole length is 15.4 meters.

EXAMPLE 14.14

Calculate the angled borehole length and angle where the bench height ($H_b$) is 14 meters (45.9 feet) with a required subdrilling of 3.3 feet and the horizontal offset off the crest to toe by survey is 11.8 feet.

Step 1 Calculate the depth of the borehole bottom from the collar as follows:

$D = H_b + l_{subdrill}$

$D = 45.9 + 3.3$

$D = 49.2$

The depth of the borehole bottom from the collar is 49.2 feet.

Step 2 Calculate the required drill angle using equation 14.10.

$A = \tan^{-1}\left(\frac{O}{D}\right)$

$A = \tan^{-1}\left(\frac{11.8}{49.2}\right)$

$A = \tan^{-1}(0.24)$

$A = 13.5$

The drill angle is 13.5°.

Step 3 Calculate the borehole length using equation 14.9.

$l = \frac{O}{\sin A}$

$l = \frac{11.8}{\sin(13.5)}$

$l = \frac{11.8}{0.233}$

$l = 50.6$

The borehole length is 50.6 feet.

Velocity Matching

Velocity matching refers to the velocity of detonation (VOD) of the explosive relative to the sonic velocity of the rock. Increasing the VOD to match the rock's sonic velocity generally improves fragmentation. Field research has shown that VODs exceeding the rock's sonic velocity by 30% does not provide additional fragmentation (Chiappetta, 1991). A rock's sonic velocity is a reflection of the speed a sound wave can travel through it. Rock sonic velocity is directly related to its elasticity and dynamic response to stress.

Additional Resources

Chiappetta, R. Frank. 1991. Generating site-specific custom blast designs with onsheet blast monitoring instrumentation. Blasting Analysis International, Inc. (BAI). 3rd High-Tech Seminar. June 2 – 7, San Diego. CA. BAI, Allentown, PA.

Dick, Richard A., Larry R. Fletcher, Dennis V. D'Andrea. 1982. Explosives and Blasting Procedures Manual. IC 8925. Twin Cities Research Center, Bureau of Mines. Minneapolis, MN.

Hustrulid, William. Rabert m and Calvin Konya. 1985. The effect of bench stiffness with increasing blasthole length. International Society of Explosives Engineers (ISEE) Research Proceedings of the 11th Annual Conference on Explosives and Blasting Technique. February 17 – February 21. San Diego CA: ISEE, Cleveland, OH.

Institute of Makers of Explosives (IME) 2010. Safety Library Publication 12, Glossary of Commercial Explosives Industry Terms. IME, Washington, DC.

Konya, Calvin J. and Edward J. Walter. 1990. Surface Blast Design. ch. 11 pp. 22- 34. Prentice Hall, Inc., Englewood Cliffs, NJ

Lilly Explosives Company. 1992. Efficient Blast Management. Charlestown, WV.

Oriard, Lewis L. 2005. Explosives Engineering, Construction Vibrations and Geotechnology. International Society of Explosives Engineers, Cleveland, OH.

Postquick, Corey. 1999. Application and measurement of emulsion energy and at work in rock. Blasting Analysis International, Inc. (BAI). Proceedings of the 1st Annual High-Tech Seminar: State-Of-the-Art Blasting Technology Instruments and Applications. May 17 – 19, San Diego CA. BAI, Allentown, PA.

Pugliese, Joseph M. 1982. Designing blast patterns using empirical formulas – A comparison of calculated patterns with plans used in quarrying limestone and dolomite, with geological considerations. IC 8892. Twin United States Bureau of Mines, Twin Cities Research Center, Minneapolis, MN.

Roy, P. Pal and Dr. S. K. Singh. 1988. New burden and spacing formulae for optimum blasting. International Society of Explosives Engineers (ISEE) Research Proceedings of the 24th Annual Conference on Explosives and Blasting Technique. February 1 – 5, New Orleans, LA. ISEE, Cleveland, OH.

Rustan, A., 1998. Rock blasting terms and symbols. Taylor and Francis, Inc., Rotterdam.

Workman, Lynn, and Peter Calder, 2003. Blast design and implementation, Part E. Calder and Workman, Inc., Bismarck, ND.

Hustrulid, William. 1999. Blasting Principles for Open Pit Mining No. 1 General Design Concepts. A.A. Balkema Publishing; Rotterdam, Netherlands.

International Society of Explosives Engineers (ISEE). 1998. ISEE Blasters' Handbook™, 17th Edition. ISEE, Cleveland, OH