RI 9691 

REPORT OF INVESTIGATIONS/2013 

A New Perimeter Control Blast Design Concept 

for Underground Metal/Nonmetal 

Drifting Applications 

Department of Health and Human Services 

Centers for Disease Control and Prevention 

National Institute for Occupational Safety and Health 
 

 

 

Report of Investigations 9691 

A NEW PERIMETER CONTROL BLAST DESIGN CONCEPT FOR 

UNDERGROUND METAL/NONMETAL DRIFTING APPLICATIONS 

Stephen R. Iverson, William A. Hustrulid, and Jeffrey C. Johnson 

DEPARTMENT OF HEALTH AND HUMAN SERVICES
 Centers for Disease Control and Prevention
 

National Institute for Occupational Safety and Health
 

Office of Mine Safety and Health Research
 

Pittsburgh, PA • Spokane, WA
	

March 2013 
This document is in the public domain and may be freely 

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DHHS (NIOSH) Publication No. 2013–129 

March 2013 
Table of Contents
 

Abstract...................................................................................................................................... 1
 

Introduction ................................................................................................................................ 2
 

The Practical Damage Radius.................................................................................................... 3
 

Background........................................................................................................................ 3
 

Buffer Holes in Practical Design ......................................................................................... 6
 

Modified Ash Energy (MAE) Approach....................................................................................... 8
 

Background........................................................................................................................ 8
 

Equation............................................................................................................................. 9
 

Calculation Examples........................................................................................................11
 

Other Damage Models Considered ...................................................................................12
 

Blast Damage Experiments.......................................................................................................14
 

Experimental Methodology ................................................................................................14
 

Concrete Block and Explosive Properties..........................................................................21
 

Micro-Velocity Probe .........................................................................................................22
 

Crosshole Velocity Probes ................................................................................................22
 

Wave Attenuation Measurements......................................................................................23
 

Visual Assessment of Radial Crack Damage.....................................................................24
 

Practical Damage Limit .....................................................................................................25
 

Wave Attenuation Results .................................................................................................25
 

Block 1 Damage Assessment............................................................................................26
 

Block 2 Damage Assessment............................................................................................29
 

Block 3 Damage Assessment............................................................................................31
 

Discussion.........................................................................................................................33
 

Buffer Hole Design Concept......................................................................................................34
 

Perimeter Charging ...........................................................................................................36
 

Discussion.........................................................................................................................37
 

Analysis of Successful Perimeter Control Designs ....................................................................38
 

Introduction .......................................................................................................................38
 

Example of a Successful String-Loading Perimeter Control Design...................................38
 

Example of a Successful Bottom Charge Perimeter Control Design..................................42
 

Example of a Successful Swedish Decoupled Cartridge Perimeter Control Design ...........44
 

Example of a Successful Canadian Controlled Blast Design .............................................48
 

Example of a Successful Spanish Railway Tunnel Construction Controlled Blast Design..50
 

Discussion.........................................................................................................................52
 
Conclusions ..............................................................................................................................53
 

Recommendations ....................................................................................................................53
 

Acknowledgments.....................................................................................................................54
 

References ...............................................................................................................................55
 

Project Bibliography ..................................................................................................................58
 

Appendix A: Perimeter Hole Covolume Wall Pressure Calculation Method Used for Successful
 

Design Examples ......................................................................................................................61
 
Figures
 

Figure 1. Diagrammatic representation of the design sectors involved in a drift round. After
 

Holmberg [1982]. ....................................................................................................... 4
 

Figure 2. Contour sector bounded by the contour row and the buffer row. ................................. 5
 

Figure 3. The charge concentrations in the holes close to the contour are adjusted so that the
 

damage zone from each hole coincides with the expected crack limit. After Holmberg 

[1982]......................................................................................................................... 5
 

Figure 4. Diagrammatic representation of the crushed, cracked, and damaged zones 

surrounding a blasthole.............................................................................................. 6
 

Figure 5. Radial crack damage circles applied to holes emphasized in the Holmberg [1982]
 

perimeter damage extent diagram. ............................................................................ 7
 

Figure 6. Practical damage circles only applied to buffer holes in the Holmberg [1982] perimeter
 

damage extent diagram. ............................................................................................ 7
 

Figure 7. Isometric view of Ash’s blast design parameters. ........................................................ 8
 

Figure 8. MAE practical damage Rd/rh plot of selected explosives for a range of rock densities.
 

.................................................................................................................................12
 

Figure 9. Engineering drawing of block 1 blasthole and strain gage sensor locations. ..............16
 

Figure 10. Engineering drawing of block 2 blasthole and strain gage sensor locations..............17
 

Figure 11. Engineering drawing of block 3 blasthole and strain gage sensor locations..............18
 

Figure 12. Block 1 prior to blasting showing flyrock barrier wall installed...................................19
 

Figure 13. Block 2 after blasting and prior to wire sawing showing remnants of blasthole 

location. ....................................................................................................................20
 

Figure 14. Block 3 during wire sawing to expose radial fractures. .............................................21
 

Figure 15. Plot of radial strain measurements for all three block experiments showing regression
 

curve fit of all data.....................................................................................................26
 

Figure 16. Block 1 MVP velocity vs. distance from blasthole. ....................................................26
 

Figure 17. Block 1 postblast inspection hole damage assessment using the crosshole sonic Pwave velocity probes.................................................................................................27
 

Figure 18. Block 1 photograph/quadrant overlay of postblast radial crack damage visible from
 

wire saw cut..............................................................................................................28
 

Figure 19. Block 1 postblast number of cracks counted with distance from the blasthole. .........29
 

Figure 20. Block 2 inspection hole damage assessment using the crosshole sonic P-wave 

velocity probes..........................................................................................................29
 

Figure 21. Block 2 photograph/quadrant overlay of postblast radial crack damage visible from
 

wire saw cut..............................................................................................................30
 

Figure 22. Block 2 number of cracks counted with distance from the blasthole. ........................31
 

Figure 23. Block 3 postblast shear wave velocity as a function of distance from the blasthole 

with sensors oriented vertically. ................................................................................31
 
Figure 24. Block 3 photograph/quadrant overlay of postblast wire saw section near blasthole 

collar showing cracks. ...............................................................................................32
 

Figure 25. Block 3 postblast number of cracks vs. distance from blasthole. ..............................33
 

Figure 26. Summary of damage limit results of block experiments compared to MAE Rd. .........34
 

Figure 27. Buffer row hole locations, buffer hole practical damage circles, and perimeter holes at
 

the cusp locations. ....................................................................................................35
 

Figure 28. Range of buffer hole spacing conditions showing percentage of area covered by
 

circles formed using the Rd damage radius. ..............................................................36
 

Figure 29. Photograph and diagrams of string-loading bulk emulsion using constant pumping
 

and variable rate extraction of the injection hose [Fauske 2003; Hustrulid 2008]. .....38
 

Figure 30. Example showing drill hole locations of a successful blast design in Australia that
 

uses string-loaded emulsion. ....................................................................................40
 

Figure 31. Example of a successful blast design in the United States with bottom-charged
 

perimeter holes. ........................................................................................................43
 

Figure 32. Example of the successful blast design in Sweden with decoupled perimeter
 

cartridges..................................................................................................................46
 

Figure 33. Example of a successful blast design in Canada with decoupled perimeter cartridges
 

(shown are only the locations of the buffer, perimeter, and lifter holes).....................48
 

Figure 34. Example of the successful railway tunnel blast design in Spain................................50
 
Tables

Table 1. Selected explosives, properties, and calculated Rd/rh damage results for a range of 

rock densities............................................................................................................11
 

Table 2. Comparison of experimental setup parameters to actual mine drifting parameters ......14
 

Table 3. Concrete block properties tested.................................................................................21
 

Table 4. Explosive properties of the emulsions used in the blast experiments...........................22
 

Table 5. PPV limits for blast damage types for hard gneiss or granite [Persson et al. 1994] .....23
 

Table 6. Estimated Persson PPV damage limits for the experimental blocks using the
 

compressive strength PPV factors [Persson et al. 1994]...........................................24
 

Table 7. Practical damage limit results for the three blocks tested ............................................25
 

Table 8. Strain measurements and PPV conversions for the three block experiments ..............25
 

Table 9. Block 1 number of radial cracks visible in wire saw cut................................................28
 

Table 10. Block 2 number of radial cracks visible in wire saw cut..............................................30
 

Table 11. Block 3 number of radial cracks visible in wire saw cut..............................................32
 

Table 12. Buffer row design concept parameters and damage compared to successful
 

Australian design ......................................................................................................41
 

Table 13. Buffer row design concept parameters and damage compared to successful U.S.
 

design.......................................................................................................................44
 

Table 14. Successful Swedish controlled blast design explosive properties ..............................45
 

Table 15. Properties of rock at the example mine .....................................................................47
 

Table 16. Buffer row design concept parameters and damage compared to successful Swedish 

design.......................................................................................................................47
 

Table 17. Explosives used for Canadian successful design ......................................................49
 

Table 18. Buffer row design concept parameters and damage compared to successful Canadian
 

design.......................................................................................................................50
 

Table 19. Buffer row design concept parameters and damage compared to successful tunnel
 

design.......................................................................................................................51
 

Table 20. Covolume spreadsheet calculation for Australian design ...........................................65
 

Table 21. Covolume spreadsheet calculation for bottom charge design....................................65
 

Table 22. Covolume spreadsheet calculation for Swedish decoupled design ............................66
 

Table 23. Covolume spreadsheet calculation for Canadian design ...........................................66
 

Table 24. Covolume spreadsheet calculation for Spanish design..............................................67
 
ACRONYMS AND ABBREVIATIONS
 

ANFO Ammonium Nitrate Fuel Oil 

CSL Crosshole sonic logging 

MAE Modified Ash Energy 

MSHA Mine Safety and Health Administration 

MVP Micro-velocity probe 

NIOSH National Institute for Occupational Safety and Health 

P-wave Compressional wave 

PETN Pentaerythritol tetranitrate 

PPV Peak particle velocity 

RBS Relative bulk strength 

RMR Rock Mass Rating 

RWS Relative weight strength 

S-wave Shear wave 

SSE Solids stabilized emulsion 

UCS Uniaxial compressive strength 

VOD Velocity of detonation 
UNIT OF MEASURE ABBREVIATIONS 

% percent 

μsec microsecond 

atm atmosphere 

cm centimeter 

cm 3 cubic centimeter 

cm3

/g cubic centimeter per gram 

oC degree Celsius 

g gram 

g/cm3 gram per cubic centimeter 

GPa gigapascal 

K Kelvin 

kcal/kg kilocalories per kilogram 

kg kilogram 

kg/m kilogram per meter 

kg/m3 kilogram per cubic meter 

km/sec kilometer per second 

L/kg liter per kilogram 

m meter 

MHz megahertz 

MJ/kg megajoules per kilogram 

mm millimeter 

MPa megapascals 

ms millisecond 

m/sec meters per second 

m 2 square meter 

m 3 cubic meter 

m3

/m cubic meter per meter 

sec second 

 

A NEW PERIMETER CONTROL BLAST DESIGN CONCEPT
 

FOR UNDERGROUND METAL/NONMETAL DRIFTING
 

APPLICATIONS
 

Stephen R. Iverson, William A. Hustrulid, and Jeffrey C. Johnson
 

Office of Mine Safety and Health Research
 

National Institute for Occupational Safety and Health
 

Abstract 

This report presents a new concept in perimeter control blasting for underground 

metal/nonmetal mine drifting applications focusing on the importance of the buffer holes in a 

blast design. The new blast design concept applies the understanding of radial damage that is 

caused by the buffer hole column charge. Buffer hole radial damage is defined by a practical 

damage limit applied to the rock lying between the buffer holes and the perimeter. A favorable 

comparison was made between five successful controlled blast designs and the concept of 

practical damage limit. This concept is easy to use and acceptance of the approach would only 

require mines to conduct trial blasts to test the design theory. 
Introduction
 

Drift driving is an important part of bringing metal/nonmetal mines into operation. Drifting 

provides access to an orebody for openings such as main haulage ways, main levels, ramps, 

crosscuts, sublevels, etc. The width, height, cross-sectional shape, and length of a drift depend on 

the use of the various mine openings. The typical method of drifting is to drill and blast. Blasting 

is an economical method to break rock, but the consequence of blasting is damage to the 

perimeter of the opening. The damage results in overbreak or rock that was not designed to be 

removed, loose rock to be scaled, and permanent damage to the remaining perimeter. Both the 

lack of a controlled blast design and of precision drilling are the main causes of perimeter 

damage. Perimeter damage was assessed in 2006 and 2007 by researchers from the National 

Institute for Occupational Safety and Health (NIOSH) during field investigations of blasting 

practices at mines in the United States [Iverson et al. 2007; Warneke et al. 2007; McHugh et al. 

2008a; McHugh et al. 2008b]. It was found that mines were drilling blastholes without precision 

drills and at best were only able to maintain parallel blasthole orientations. Further, the blast 

designs were based primarily on miners’ experience and capabilities. The results were a wide 

range of blast patterns, cut designs, hole spacings for the perimeters, the number of empty holes 

and spacing for line drilling, perimeter burdens, explosive types used, use of perimeter hole 

decoupling to prevent damage, blast round length, and variability in cross-sectional arch design 

and implementation. 

The field investigations into drifting practices were important to outline safety problems 

linked to ground control. The obvious safety problems are: 

	 Overbreak results in wider spans that require additional ground support and an increased 

likelihood of failure if not properly assessed. 

	 Rough and undulating back and wall surfaces occur due to aggressive blasting and likely 

increase the hazards associated with scaling and the installation of bolts and support 

accessories. 

	 Lack of perimeter control by aggressive blasting will damage the perimeter to the point 

where more scaling is required and more potential loose rock could develop. 

	 Flat-arched backs impose additional bolting requirements where a rounded arch will 

typically aid in supporting the back. 

Mine Safety and Health Administration (MSHA) accident statistics were reviewed by 

NIOSH engineers in 2006 and it was determined that fall-of-ground accidents in metal and 

nonmetal mines could be further prevented by improving the blasting methods. Interaction with 

the mining community to identify critical hazards was completed through field investigations. 

Blasting has a tremendous influence on roof and rib stability [Iannacchione and Prosser 

1997]. Precision drilling technology and controlled blast designs are available to accomplish 

better blasting perimeter control. The latest technology in drill jumbos are computer controlled 

and can drill precision holes using an engineered design. Mines are beginning to use these drills, 

and the future will hold the results as to the benefits of these technologies and methods. The blast 

designs currently used at the majority of mines in the United States are less developed and are 

based on dated design methods. Current controlled blast designs are based on closely spaced 

2
 
perimeter holes using the blasthole diameter to determine the perimeter spacing and burden and 

are not based on the explosive quantity or type for the perimeter holes and do not consider the 

effect of perimeter damage by the buffer holes [Holmberg 1982; Konya 2006]. 

This report proposes for the mining community an easy-to-use blast design method that 

includes improvements for determining the perimeter burden based on the effect of damage from 

buffer holes. This means that the distance or burden between the perimeter holes and the next 

line of blastholes defined as the buffer holes is determined by the damage caused by the buffer 

hole detonations. The new design method also includes the concept of locating the perimeter 

holes in an alternating arrangement in relation to the buffer hole locations. 

The research and development of the new design described in this report was aimed at: 

	 Identifying the effectiveness of perimeter control in current drift designs 

	 Conducting experiments to determine the blast damage extent and factors influencing 

damage 

	 Studying blast damage models 

	 Packaging the blast damage models into an engineer and miner friendly design 

concept 

This report includes: 

	 An introduction to the new perimeter control design philosophy and concept 

	 A blast damage model for calculating practical damage 

	 An evaluation of single hole blasts conducted in large concrete blocks to understand 

radial crack damage 

	 A description of five successful design examples demonstrating application of the 

design concept 

Simple assumptions (that the rock mass is homogenous, isotropic, and a structure-free rock 

mass) are made in regards to drift design to focus on the importance of the practical damage limit 

and buffer row placement. Future research will be needed to address rock structure, detonation 

timing and sequencing, the interaction between adjacent holes detonated in a design, and the 

potential for improvements in fragmentation and perimeter control using electric detonators. 

The Practical Damage Radius 

Background 

Perimeter control blasting is commonly applied to the roof but increasingly it is being applied 

to the walls as well. Figure 1 illustrates the different design sectors as distinguished by Holmberg 

[1982] for a drift round. 

3
 
 .1 25EB 

 

Figure 1. Diagrammatic representation of the design sectors involved 

in a drift round. After Holmberg [1982]. 

To better understand the traditional approach to perimeter control, the following is a 

description of the Holmberg [1982] algorithm. One begins by designing the cut, the contour, and 

the lifter sectors. When the particular designs for each of the sectors have been completed, they 

are added to the overall drift design. The so-called stoping holes are then added as needed. The 

implication is that the holes/explosives included in the different sectors are responsible for 

removing the associated rock. In the example shown in Figure 2 the width (“burden”) of the

contour sector is denoted as B. Particularly in hard, strong rock it is expected that the amount of 

explosive charge required to remove the contour sector of rock is quite high, and this 

subsequently would place quite strong requirements on the perimeter row design. For the 

Holmberg design, the contour (perimeter) burden is simply determined based on the spacing of 

the perimeter holes using: 

 (1)

where	 B = perimeter burden (m), 

E = perimeter hole spacing (m),
 

and 

E  kd

(2) 

where	 k = 15 to 16, and
 

d = perimeter blasthole diameter (m).
 

4
 
Holmberg [1982] suggests analysis of the stoping, buffer, and perimeter holes for perimeter 

damage using critical peak particle velocity (PPV). The damage extent is illustrated in Figure 3. 

Figure 2. Contour sector bounded by the contour row and the buffer row. 

Figure 3. The charge concentrations in the holes close to the contour are adjusted so that the 

damage zone from each hole coincides with the expected crack limit. After Holmberg [1982]. 

Figure 3 shows potential contributions of the perimeter, buffer holes, and stoping holes in 

breaking the contour sector of rock and eventually damaging the perimeter rock. Not 

surprisingly, these contributions precondition the rock towards the perimeter. This is supported 

by Tesarik et al. [2011] who found through analysis of peak particle velocities that blast damage 

occurs beyond the subsequent row to be detonated. Preconditioning is a common phenomenon 

where blasting fragmentation results in weakening the fragments by microcracking [McCarter, 

1996]. 

5
 
 

Buffer Holes in Practical Design 

Buffer holes have previously been suggested for use in blast design as described by Hustrulid 

and Johnson [2008]. A practical damage radius (Rd) is determined for each blasthole/explosive 

combination. The damage radius calculated from the buffer holes would determine the perimeter 

burden. By “practical,” it is meant that if the rock mass lying outside of this ring were removed, 

the rock remaining within the ring would easily break apart. As can be seen in Figure 4, the 

practical damage zone consists of both crushing and cracking components. 

Figure 4. Diagrammatic representation of the crushed, cracked, and 

damaged zones surrounding a blasthole. 

The idea is to design 

the buffer holes so that their associated damage radius extends to the desired drift perimeter 

where the practical damage radius: 

d  BR

(3) 

where	 Rd = practical damage radius, and
 

B = contour row burden.
 

It is logical to consider the extent of the damage surrounding a cylindrical charge to be a 

function of the available explosive energy, or the pressure applied to the wall of the blasthole. In 

reality, it is most probably a combination of both. In reviewing Figure 4, three different limits 

can be considered: 

 crushing 

 cracking 

 practical damage 

6
 
Holmberg and Persson [1979] offered a practical approach in determining the cracking limit 

through their well-known design curves relating the PPV, radius, and the explosive charge 

concentration. Referring back to Figure 3, the graphic visually presents Holmberg and Persson’s

damage concept that the crack radii from the buffer and stoping rows of holes should not exceed 

that associated with the contour holes. When circles are drawn in the figure rather than simply 

arcs, one can easily visualize that the representation is consistent with the proposed practical 

damage radius approach to design (Figure 5).

Figure 5. Radial crack damage circles applied to holes emphasized in the 

Holmberg [1982] perimeter damage extent diagram.
 

 It must be emphasized, however, that the PPV 

approach is based on the cracking radius rather than on the practical damage radius. Figure 6 

provides a visual representation of the smaller practical damage limit applied to the buffer holes 

for the same diagram. 

Figure 6. Practical damage circles only applied to buffer holes in the
 

Holmberg [1982] perimeter damage extent diagram.
 

7
 
 

In some cases, for example in tunnel construction, the cracking radius is of prime concern 

and must be carefully controlled. In most cases, however, the primary control is on the size and 

shape of the excavation as denoted by the excavation limits. In design, one then utilizes, first and 

foremost, the practical damage limit. One can estimate the expected cracking limits but this is 

done more for interest than as a control. Through the use of presplitting of the perimeter row one 

can, at least in principle, control both the cracking limits and the excavation limits. 

Modified Ash Energy (MAE) Approach 

Background 

In 1963, Ash [1963a,b,c,d] published his now-classic papers dealing with blast design in 

open pit mines and quarries. Figure 7 below is an isometric representation of two blastholes from 

a typical open pit mine or quarry and shows Ash’s various design parameters. 

Figure 7. Isometric view of Ash’s blast design parameters [Ash 1963a,b,c,d]. 

Using field data collected from a large number of operations, Ash [1963a,b,c,d] summarized 

the different design parameters. If these design parameters are applied to underground blast 

design, these same parameters should apply. He found that the subdrill (J), the stemming (T), the 

spacing (S) and the bench height, and the expected excavation length (H) could all be related to 

the burden (B). Most importantly, he found that for fully charged holes the burden (B) was 

related to the blasthole diameter (De) as shown in the following equation: 

8
 
DKB eB



(4)
 

 

 2RB d

where	 B = burden, 

KB = constant, and 

De = hole diameter. 

The value of KB can vary with both the rock and the explosive, though Ash recommends that 

for average rock conditions KB = 30 except when using a lighter density charge of field-mixed 

aluminum nitrate fuel oil (ANFO) where KB is reduced to a range of 20 to 25. Hustrulid [1999a] 

3 found that KB = 25 when using ANFO at a density of 0.80 g/cm in rock of medium density (2.65 

3 g/cm ). 

Ash’s work has very successfully captured the experience in blasting. The ratios are largely 

followed today with the possible exception of the hole spacing relationship. Further, the range of 

rock types, explosive types, and blasthole diameters examined by Ash provide a satisfactory 

empirical analysis for conditions and ranges also found in underground mines. Ash contends that 

KB can be modified using explosive density and rock density to values as small as 20 and as high 

as 40. This range of KB would be the limitation of its use for determining burden. Ash suggests a 

practical range for both explosive densities and rock densities. [Ash 1963a,b,c,d]. 

Equation 

Hustrulid [1999a] proposed a design procedure based on energy coverage. It was assumed 

that the holes were charged fully coupled. The radius of the damage circle used in this approach 

was obtained by equating available explosive energy to that required to produce acceptable 

3 fragmentation. In this regard, it was assumed that the use of ANFO with density 0.85 g/cm to 

3 blast an average rock of density 2.65 g/cm yielded satisfactory results when KB = 25. The 

formula expressing the ratio KB for other rock-explosive combinations becomes: 

ANFO ANFO r

ANFOe

B

s

s

K

 

 .2 65 25

1



(5) 

3 where	 ρe = density of the explosive used (g/cm ),
 

sANFO = weight strength of the explosive relative to ANFO,
 3

ρANFO = density of ANFO = 0.85 g/cm ,
 

s1ANFO = weight strength of ANFO relative to ANFO = 1, and
 3

ρr = density of the rock mass (g/cm ).
 

This formula described originally in Hustrulid and Johnson [2008] provides a value of KB 

which follows the Ash [1963a,b,c,d] and Hustrulid [1999a] recommendations regarding the 

effect of explosive energy and rock density. Based simply on the geometry of just-touching 

damage circles, 

9
 

(6) 
hBeBd 2  2rKDKRB 

 

 

hBdhBd  22 rKR or  rKR 

 

 

 

The basic Ash [1963a,b,c,d] burden formula becomes 

(7)
 

where	 rh = hole radius, and
 

Rd = damage radius.
 

From Equations 6 and 7, it follows that 

(8) 

Finally, one obtains the relationship 

ANFO r

ANFOe

h

d

s

r

R



 .2 65

 25

(9) 

It is recognized that 

RBS s

ANFO

ANFOe 





(10) 

where RBS = bulk strength relative to ANFO. 

The relative bulk strength (RBS) is often provided by explosive suppliers. Thus Equation 10 

becomes 

h r

d RBS

r

R



.2 65

 25

(11) 

As can be seen, the approach is quite simple involving available explosive properties and the 

density of the rock as inputs. Thus, this approach is attractive for mine application. 

10
 
Calculation Examples 

An explosive properties dataset of 15 types was evaluated as applied to a 48-mm-diameter 

blasthole. The dataset is shown in Table 1 and plotted in Figure 8. 

11
 

Table 1. Selected explosives, properties, and calculated Rd/rh damage results 

for a range of rock densities 

Explosive RBS 2.2 2.5 2.7 2.9 3.2 

Orica Titan SSE* 0.66 22.3 21.0 20.2 19.5 18.5 

Orica Titan SSE 0.70 23.0 21.6 20.8 20.0 19.1 

Orica Titan SSE 0.74 23.6 22.2 21.3 20.6 19.6 

Orica Titan SSE 0.78 24.2 22.7 21.9 21.1 20.1 

Orica Titan SSE 0.82 24.8 23.3 22.4 21.6 20.6 

Orica Titan SSE 0.86 25.4 23.8 22.9 22.1 21.1 

Orica Titan SSE 0.90 26.0 24.4 23.5 22.6 21.5 

DYNO MIX 0.92 26.3 24.6 23.7 22.9 21.8 

Orica Titan SSE 0.94 26.5 24.9 24.0 23.1 22.0 

DYNO Titan 7000 0.94 26.5 24.9 24.0 23.1 22.0 

DYNO BlastEX 1.06 28.2 26.5 25.5 24.6 23.4 

Orica Amex 1.10 28.8 27.0 26.0 25.1 23.9 

DYNO BlastEX Plus 1.15 29.4 27.6 26.5 25.6 24.4 

DYNO AP (tamped) 1.24 30.6 28.7 27.6 26.6 25.3 

Orica Senatel Magnafrac 

(tamped) 

1.20 30.1 28.2 27.1 26.2 24.9 

* Solids stabilized emulsion 

Ash suggests his KB value can 

3 3 be modified based on a range of rock densities from 2.2 g/cm to 3.2 g/cm . A rock density of 2.7 

3 g/cm was considered most common. Lighter charges will have a lower KB while heavier charges 

will have a larger KB. This is equally true with the Rd/rh ratio in Figure 8 where the lighter 

charges are in the range of 25 and the heavier charges are closer to 30. A range of rock densities 

3 3 from 2.2 g/cm to 3.2 g/cm as described by Ash was applied, and the rock densities are shown 

as separate curves in Figure 8. [Ash 1963a,b,c,d]. 
Figure 8. MAE practical damage Rd/rh plot of selected explosives for a range of rock densities. 

Other Damage Models Considered 

Other damage methods for assigning a damage radius were investigated. These methods 

ranged from those that are empirically based to some with very solid theoretical roots. In theory 

or through support of other empirical datasets, some of the other damage models mentioned 

could be viable choices for determining the practical damage limit. The authors of this paper 

contend that the Modified Ash Energy (MAE) approach has the most comprehensive and 

empirically supported dataset and is based on commonly used explosives and rock properties. If 

one were to choose an approach for a first approximation of practical damage, it would be the 

MAE approach. 

Short discussions on other damage models investigated for use in determining the practical 

damage radius are listed below: 

	 The Holmberg-Persson approach [Holmberg and Persson 1979] has been widely 

accepted and applied. In addition, by serving as a model for the development of other 

approaches, it has greatly contributed to advancing the field. 

	 The rock constant approach is based on work by Persson et al. [1994] where the rock 

3 constant c is the amount of explosive (kg/m ) needed for breaking loose the rock at 

the toe in a defined blasting geometry. The rock constant ranges from 0.2 to 0.6. The 

rock constant equation for burden calculation can be found in Holmberg [1982] and 

in Persson et al. [1994]. 

	 The modified Ash pressure model, as described in Hustrulid and Johnson [2008] and 

similar to the MAE model, was modified from work by Ash [1963a,b,c,d] and 

Hustrulid [1999a]. The RBS in the MAE approach is replaced with the explosion 

pressure. The explosion pressure is derived from an equation using explosive density 

and velocity of detonation (VOD). 

12
 
	 The Sher pressure-based approach to calculating blast damage is dependent on the 

explosive properties for calculating the explosion pressure and the calculation of 

damage based on the strength and elasticity of the rock mass [Sher 1997; Sher and 

Aleksandrova 1997; Sher and Aleksandrova 2007]. 

	 The Drukovanyi approach is based on Drukovanyi et al. [1976] who presented a 

calculation for zones of fine crushing and radial fissures for exploding cylindrical 

charges. 

	 The Johnson model is based on the Hustrulid Bar test as described by Johnson [2010]. 

In this model, it was assumed that the damage radius was directly related to the 

damage done by the shock wave moving through the rock mass. The gas pressure was 

not considered. 

	 The Etkin approach is based on Etkin et al. [2001] that describes a calculation method 

for explosives selection and placement to protect the perimeter from damage. The 

Etkin approach is empirical in that a large amount of field data was collected to 

derive a blasting resistance factor and determination based on rock classification, 

joint spacing, and joint width. The Etkin approach is promising because it takes into 

account the rock structure. 

	 The Favreau approach is based on work by Favreau [1969] who presented a 

mathematical solution for the prediction of strain waves generated when a fully 

coupled explosive detonates inside a spherical cavity in an infinite, isotropic, 

homogeneous medium. These passing strain waves produce particle vibrations. The 

difficulty is in converting the spherical equation to a column charge equation. This 

was attempted using a string of elemental spheres that represent a column charge 

similar to the Holmberg-Persson approach. 

	 The Neiman approach pertains to a cylindrical charge as originally presented by 

Neiman [1979] but modified to take into account actual rock behavior and various 

explosives. The hydrodynamic approach was developed to predict the extent of blast 

damage caused by an explosive charge in an idealized rock mass as characterized by 

the peak particle velocity (PPV). A calibration factor was introduced to account for 

the actual behavior of the explosive in a particular rock mass. This factor is obtained 

by conducting a field test in which particle velocity measurements are made as a 

function of distance away from the charge [Tesarik and Hustrulid 2009]. 

13
 
Blast Damage Experiments
 

Blasting experiments were conducted in large concrete blocks that simulate rock [Iverson et 

al. 2009; Johnson 2010]. Results from these experiments were compared to the MAE practical 

damage limit Rd. 

Experimental Methodology 

The experimental method for assessing the damage level of the MAE Rd calculation is to 

have blocks with a range of rock densities and explosive RBS values to properly assess the range 

of rock densities and RBS values shown in Figure 8. The experiments that were conducted were 

limited to two types of emulsion and a narrow range of simulated rock densities. The damage 

results from the experiments could indicate a definite change in damage at the MAE Rd limit or 

at least provide better understanding of and quantify the damage at the Rd limit. 

The experiments were explosive detonations in large concrete blocks and a blasthole of fullscale diameter, length, and radial extent. Full-scale tests best represent actual mine drifting 

conditions. The tests were comparable to actual drifting dimensions, explosives, and rock 

densities. A comparison of actual to experimental parameters is shown in Table 2. 

Table 2. Comparison of experimental setup parameters to actual mine drifting parameters 

Parameters Mine drifting * Block experiment 

Diameter (mm) 

Length (m) 

Radial extent (m) 

Rock density (kg/m3

) 

RBS 

44–48 

2.4–3.7 

na†

2,200–3,200 

0.6–1.2 

38 

1.2–1.8 

2.6–3.6 

2,100–2,310 

1.20–1.24 

* Typical ranges of dimensions were found in field investigations and MAE input properties 

from Ash [1963a,b,c,d]. 

†na = not applicable 

The explosive for the experiments was packaged emulsion. Emulsion provides uniformity in 

the explosive as compared to ANFO where the amount of pressure when pneumatically loading 

can result in varied in-place properties. The use of ANFO would also have required a larger hole 

diameter. The packaged emulsion was tamped to full coupling with the blasthole wall. 

Damage was measured using various techniques including: (1) wire sawing to expose and 

measure the crack count and length, (2) seismic probes using inspection holes to determine the 

combined effect of microcrack damage and radial crack damage, and (3) measured strain 

applying the peak particle velocity (PPV) to relate the Holmberg-Persson [1979] damage limits 

to distance. 

14
 
The analysis of the blocks included comparing the measured damage limits for each test to 

the MAE Rd for the specific simulated rock density and explosive properties. Finally, damage at 

the calculated MAE Rd practical damage limit was assessed. 

The experimental design setups for blocks 1, 2, and 3 are shown in Figure 9, Figure 10, and 

Figure 11, respectively. The blastholes were 38 mm in diameter and drilled parallel to the free 

surface of each block. Three different burden dimensions and two brands of emulsion explosives 

were employed. The charges were fully coupled. The waves generated by the detonation process 

were measured at several distances away from the charge axis. The radial strain measurements 

collected from the embedded gages during the blast tests were converted to PPV. These were 

used to calculate PPV-distance curves to be related to the damage. The resulting radial cracks 

were observed on wire-sawn surface cuts made perpendicular to the charge column orientation. 

Postblast inspection holes were drilled into the damaged blocks. In-hole S-wave and crosshole Pwave velocities were measured. These measurements were used to determine if damage varied 

with distance from the charge. Preblast core samples were tested for static physical properties. 

15
 
16
 

Figure 9. Engineering drawing of block 1 blasthole and strain gage sensor locations. 
Figure 10. Engineering drawing of block 2 blasthole and strain gage sensor locations. 

17
 
Figure 11. Engineering drawing of block 3 blasthole and strain gage sensor locations. 

Test blocks 1 and 2 were 3.7 m long, 2.4 m wide, and 1.8 m high. Test block 3 was 3.0 m 

long by 3.0 m wide by 1.5 m high. All of the blastholes were drilled horizontally at mid-height 

and parallel to the short side of the block. Block 1 had a burden of 15 cm, block 2 had a burden 

of 23 cm, and block 3 had a burden of 46 cm. 

In test blocks 1 and 2 the blastholes exited through the back of the block. Both ends of the 

blastholes were stemmed with 30 cm of clay. The resulting charged length was 1.8 m. In test 

block 3 the blasthole was drilled to a depth of 1.8 m and charged over a length of 1.2 m. The 

charge was not stemmed. Test block 1 as shown in the photograph in Figure 12 is prior to the 

blast experiment and shows the drilled blasthole and some of the fly rock barrier walls. Test 

block 2 as shown in the photograph in Figure 13 is after the blast experiment but before wire 

sawing to further expose the radial fracturing. Test block 3 as shown in a photograph in Figure 

14 is during wire sawing to expose the radial fractures. Note that the grout columns are exposed 

from the access holes where the strain gages were located. 

18
 
Figure 12. Block 1 prior to blasting showing flyrock barrier wall installed. 

19
 
Figure 13. Block 2 after blasting and prior to wire sawing showing remnants of blasthole location. 

20
 
Figure 14. Block 3 during wire sawing to expose radial fractures. 

Concrete Block and Explosive Properties 

The concrete block properties as determined from cores extracted from the blocks are listed 

in Table 3. 

Table 3. Concrete block properties tested 

Property Test 

block 1 

Test 

block 2 

Test 

block 3 

3 Density (kg/m ) 

Unconfined compressive strength (MPa) 

Brazilian tensile strength (MPa) 

Young’s modulus (MPa)

Speed of sound (m/sec) 

2,270 

42 

2.8 

29,500 

3,800 

2,310 

44 

3.1 

32,000 

3,930 

2,100 

21 

na* 

13,100 

2,640 

*na = not applicable 

21
 
The Orica brand Senatel Magnafrac explosive was used in test blocks 1 and 2. The Dyno 

Nobel brand Dyno AP was the explosive used in test block 3 (block 3 was tested first and then 

followed by block 1 and block 2). The reason for changing to Senatel Magnafrac was that the 

Dyno AP explosive was no longer sold in small quantities. These stick explosives were 

considered fully coupled because tamping was performed after each stick was inserted (the first 

stick containing the detonator was not tamped). The published properties for these explosives are 

listed in Table 4. 

Table 4. Explosive properties of the emulsions used in the blast experiments 

Explosive Property Dyno AP Senatel Magnafrac 

3 Density (kg/m ) 

Relative weight strength (RWS) 

Relative bulk strength (RBS) 

1,150 

0.88 

1.24 

1,110 

0.91 

1.20 

Micro-Velocity Probe 

After blasting, 76-mm-diameter boreholes were collared and drilled into the blocks at one or 

more locations along the remaining half-barrel of the blastholes. The walls of these inspection 

holes were smooth and provided a good surface for making measurements using the MicroVelocity Probe (MVP). This probe, in the form of a shuttle, is an S-wave generator and receiver. 

In this case, transmit-and-receive shoes were spaced at 10 cm. The first measurement was made 

with the MVP positioned just inside the collar and measurements were made at intervals of 2.5 

cm. By observing the variation of S-wave velocity with depth, one obtains an indication of the 

extent of damage. 

Crosshole Velocity Probes 

The Crosshole Sonic Logging (CSL) testing method was used to evaluate the condition of the 

damaged block between pairs of parallel inspection holes. The inspection holes were collared on 

the remaining half-barrels and drilled horizontally and perpendicular. The CSL test involves 

filling the holes with water, inserting a pair of hydrophones into the parallel holes, and then 

simultaneously withdrawing the hydrophones by cables attached to a distance-measuring wheel 

while ultrasonic P-wave pulses are sent from the source hydrophone to the receiver hydrophone. 

The collars of the horizontal inspection holes were extended using plastic sleeves to provide a 

dam for maintaining water in the holes while testing. The system was set to send pulses every 1.3 

cm as the hydrophones were pulled out. The test can be conducted quickly and the results easily 

analyzed. The CSL results are displayed as a plot of P-wave velocity versus distance into the 

boreholes. A decrease in the P-wave velocity of the block material between the boreholes near 

the blasthole is an indication of damage. 

22
 
Wave Attenuation Measurements 

A line of instrumentation holes oriented perpendicular to the charge axis was drilled from the 

top surface of each block for the purpose of monitoring the strain waves generated by the blast. 

The hole depth extended to the blasthole elevation. A specially prepared synthetic rock 

transducer to which four strain gages had been mounted (3 horizontal and 1 vertical) was then 

grouted at the bottom of each hole. The grout was of similar density and elasticity compared to 

the blocks, and strain waves were expected to pass through the grouted sensor locations. Of the 

three horizontal gages, one was mounted parallel to the charge axis, one perpendicular, and one 

at an angle of 45°. Data were collected at 1 MHz. Peak strain data from the horizontal gages that 

oriented perpendicular to the charge axis were plotted (Figure 15). The other strain orientations 

were not used in the presented damage models, thus they are not reported. From the regression 

curve fit, the radial strain values at a given distance are converted to peak particle velocity 

(PPV). The PPV values as presented by Persson et al. [1994] as they relate to damage were 

compared to the results. The damage limits and associated PPV values from Persson et al. are 

shown in Table 5 and include the Brazilian tensile strength limit from Johnson [2010]. Johnson 

found that the tensile strength of the Swedish Vanga granite was 20 MPa as tested using the 

splitting tensile strength method or twice the direct pull test value from Persson et al. [1994]. 

Table 5. PPV limits for blast damage types for hard gneiss or granite [Persson et al. 1994] 

Damage type Stress 

(MPa) 

PPV 

(m/sec) 

PPV factor relative 

to crushing 

Crushing and compressive strength limit 

(150 MPa) 

Good fragmentation 

Fragmentation limit 

Brazilian tensile strength limit (20 MPa)* 

Incipient damage and direct pull tensile 

strength limit (10 MPa) 

Incipient swelling 

150 

50 

25 

20 

10 

7 

15 

5 

2.5 

2.0 

1.0 

0.7 

1 

0.33 

0.17 

0.13 

0.07 

0.05 

*Johnson [2010] 

In Persson et al. [1994] the range of PPV damage values corresponds to a range of stress 

levels from 7 to 150 MPa. However, a factor is needed to compare the PPV damage limits to the 

much weaker experimental blocks. The PPV factors shown in Table 5 are proportional to the 

crushing PPV limit of 15 m/sec. 

Now for comparison, the PPV damage limits are determined for the synthetic rocks of 

experimental blocks 1, 2, and 3. Table 6 shows the calculated crushing PPV damage limits and 

the factored estimates for the other damage limits for each block. The crushing PPV damage 

limits were calculated using the one-dimensional relationship for PPV and each block’s 

compressive strength, speed of sound, and Young’s modulus.

23
 
Table 6. Estimated Persson PPV damage limits for the experimental blocks using the 

compressive strength PPV factors [Persson et al. 1994] 

Damage type Block 1 PPV 

(m/sec) 

Block 2 PPV 

(m/sec) 

Block 3 PPV 

(m/sec) 

Crushing 

Good fragmentation 

Fragmentation limit 

Brazilian tensile strength limit 

Incipient damage 

Incipient swelling 

5.4 

1.8 

0.92 

0.70 

0.38 

0.27 

5.2 

1.7 

0.90 

0.68 

0.36 

0.26 

4.5 

1.5 

0.77 

0.58 

0.32 

0.23 

The strain measurements will be plotted as strain vs. distance, and radial measurements of the 

estimated Persson PPV damage limits will be determined from the plot. 

Additionally, material specimens were extracted from the experimental blocks to determine 

the dynamic compressive strength based on work by Johnson [2010] using the Hustrulid Bar 

experiment. The seismic limit testing on block 3 is described in Johnson [2010], and a similar 

test was later conducted on block 1. The bar test results from block 1 were a seismic limit strain 

of 0.002 and a dynamic compressive strength of 59 MPa. The expected PPV damage limit was 

determined to be 7.6 m/sec based on a speed of sound velocity of 3,800 m/sec. The bar test 

results from block 3 indicated a seismic limit strain of 0.0016 and a dynamic compressive 

strength of 21 MPa [Johnson 2010]. The expected PPV damage limit was determined to be at a 

PPV of 4.5 m/sec. The PPV was determined using a sound velocity of 2,800 m/sec. The seismic 

limit distance is determined from a plot of measured strain converted to PPV vs. distance and 

compared to the MAE practical damage limit. 

Visual Assessment of Radial Crack Damage 

Wire sawing normal to the charge axis was used to expose sections of the damaged block for 

visual inspection. The number of cracks was counted in a quadrant from an aperture of 90° at 25

cm intervals radial from the blasthole. The lower quadrant opposite to the burden side was used 

for this crack count measure. The number of radial cracks or the number of cracks per arc length 

were compared to the MAE practical damage limit. 

-

24
 
.2 078 .0 00004 

 x radial  

Practical Damage Limit 

The practical damage limits were determined for the three concrete blocks using the MAE Rd 

approach (Table 7). 

Table 7. Practical damage limit results for the three blocks tested 

Block Rd/rh MAE 

1 30.0 0.57 

2 29.8 0.57 

3 30.8 0.58 

Wave Attenuation Results 

The radial strain measurements from all three blocks are listed in Table 8 and plotted in 

Figure 15. The seismic velocities of the waves converting strain to PPV are included in Table 8. 

The attenuation was similar for each experiment as shown by the plot and curve fit in Figure 15. 

The regression fit is exponential and was used to calculate radial limits at the uniaxial 

compressive strength (UCS) PPV limit from Persson et al. [1994], the experimental PPV seismic 

limit from Johnson [2010], and PPV values at the calculated MAE Rd limit. The equation fit is: 

(12) 

where	 εradial = the radial strain, and
 

x = distance (m).
 

Table 8. Strain measurements and PPV conversions for the three block experiments 

Distance 

(m) 

Block 1 

strain 

Block 1 

PPV 

(m/sec) 

Block 2 

strain 

Block 2 

PPV 

(m/sec) 

Block 3 

strain 

Block 3 

PPV 

(m/sec) 

0.15 0.0165 62.7 0.0195 76.635 nt nt 

0.23 nt* nt nt nt 0.007 18.48 

0.30 0.0066 25.08 0.0058 22.794 nt nt 

0.46 0.0026 9.88 0.003 11.79 0.002 5.28 

0.61 0.0013 4.94 0.0007 2.751 nt nt 

0.69 nt nt nt nt 0.001 2.64 

0.91 nt nt nt nt 0.0005 1.32 

1.14 nt nt nt nt 0.00025 0.66 

*nt = not tested 

25
 
Figure 15. Plot of radial strain measurements for all three block experiments showing 

regression curve fit of all data. 

Block 1 Damage Assessment 

The damage measurement results are shown in three figures in this section. The damage 

measurements presented are shear wave seismic MVP (Figure 16), crosshole seismic (Figure 17), 

and radial crack count (Figure 19). 

Figure 16. Block 1 MVP velocity vs. distance from blasthole. 

The S-wave velocity data in Figure 16 shows damage closer than 0.4 m from the blasthole. 

There was a drop in velocity at about 1 m where a vertical crack was found. A curve was fit to 

the data to determine the distance at which the S-wave velocity reaches background levels for the 

undamaged block. 

26
 
kaeay )(

bx





 

An exponential function was used to mathematically determine the limit of damage. The 

function was applied using the method of selected points to the data as shown in Figure 16. The 

function is described as: 

(13) 

where a = asymptote (m/sec), 

b = decay, 

x = radial distance (m), and 

k = intercept at x = 0 (m/sec). 

The selected fit for Equation 13 is a = 1,620 m/sec, b = 7, and k = 0 m/sec. The limit of 0.56-

m distance is determined at 98% of the asymptote or 1,588 m/sec. 

The seismic limit of 7.6 m/sec determined by Johnson [2010] for block 1 corresponds to 

0.46-m radial distance using the seismic velocity of 3,800 m/sec and Equation 12. The crushing 

limit of 5.4 m/sec determined using the factor for the Persson et al. [1994] crushing limit for 

block 1 corresponds to 0.54-m radial distance using the seismic velocity of 3,800 m/sec and 

Equation 12. 

Figure 17. Block 1 postblast inspection hole damage assessment using the crosshole sonic 

P-wave velocity probes.
 

Crosshole P-wave sonic logging data were collected from inspection hole pairs spaced at 

0.58 m. The results are plotted in Figure 17. The exponential function (Equation 13) was used to 

assess the limit of damage. The selected fit is a = 3,700 m/sec, b = 6, and k = 0 m/sec. The limit 

of 0.65-m distance is determined at 98% of the asymptote or 3,626 m/sec. 

27
 
 

Radial cracks in block 1 were counted per unit arc length at radial intervals from the 

blasthole (Figure 18). 

Figure 18. Block 1 photograph/quadrant overlay of postblast radial crack damage 

visible from wire saw cut. 

The number of cracks and arc angles are listed in Table 9.

Table 9. Block 1 number of radial cracks visible in wire saw cut 

Quadrant Radial 

distance 

(m) 

Number of 

cracks 

Aperture 

(degrees) 

Cracks per m arc 

length 

1 

2 

3 

4 

5 

0.25 

0.50 

0.75 

1.00 

1.25 

9 

11 

9 

5 

4 

90 

90 

90 

65 

40 

22.9 

14.0 

7.6 

4.4 

4.6 

28
 

There is a 

decrease in number of cracks per unit arc length with distance as shown in the plot in Figure 19. 
Figure 19. Block 1 postblast number of cracks counted with distance from the blasthole. 

The number of cracks at the MAE Rd limit of 0.57 m using the regression equation in Figure 

19 is 12.3 cracks per meter of arc length. The PPV as measured using strain at the MAE Rd limit 

is 4.9 m/sec. 

Block 2 Damage Assessment 

The damage measurement results are shown in the following figures. Damage measurements 

presented are crosshole seismic (Figure 20) and radial crack count (Figure 22). 

Figure 20. Block 2 inspection hole damage assessment using the crosshole sonic 

P-wave velocity probes.
 

Crosshole P-wave sonic logging data were collected from inspection hole pairs spaced at 

0.58 m. The results are plotted in Figure 20. The exponential function was used to assess the 

limit of damage. The selected fit for Equation 13 is a = 3,700 m/sec, b = 6, and k = 0 m/sec. The 

limit of 0.65-m distance is determined at 98% of the asymptote or 3,626 m/sec. 

29
 
 

The crushing limit of 5.2 m/sec that was determined using the factor for the Persson et al. 

[1994] crushing limit for block 2 corresponds to 0.56-m radial distance using the seismic 

velocity of 3,930 m/sec and Equation 12. 

Radial cracks in block 2 were counted per unit arc length at radial intervals from the 

blasthole (Figure 21).

Figure 21. Block 2 photograph/quadrant overlay of postblast radial crack damage 

visible from wire saw cut. 

Table 10. Block 2 number of radial cracks visible in wire saw cut 

Distance 

(m) 

Number of cracks Aperture 

(degrees) 

Cracks per m arc 

length 

0.25 

0.50 

0.75 

1.00 

1.25 

6 

8 

11 

6 

2 

87 

80 

90 

62 

45 

15.8 

11.5 

9.3 

5.5 

2.0 

30
 

 The number of cracks and arc angles are listed in Table 10. There is a 

decrease in number of cracks per unit arc length with distance as shown in the plot in Figure 22. 
Figure 22. Block 2 number of cracks counted with distance from the blasthole. 

The number of cracks at the MAE Rd limit of 0.57 m using the regression equation in Figure 

22 is 11.2 cracks per meter of arc length. The PPV as measured using strain at the MAE Rd limit 

is 5.0 m/sec. 

Block 3 Damage Assessment 

The damage measurement results are shown in the following figures. Damage measurements 

presented are shear wave seismic MVP (Figure 23) and radial crack count (Figure 25). 

Figure 23. Block 3 postblast shear wave velocity as a function of distance from the blasthole with 

sensors oriented vertically. 

Figure 23 shows the MVP results. The exponential function was used to assess the MVP 

measured damage limit. The selected fit as shown in Figure 23 using Equation 13 is a = 1,850 

m/sec, b = 6, and k = 0 m/sec. The limit of 0.65-m distance is determined at 98% of the 

asymptote or 1,813 m/sec. 

31
 
The seismic limit of 4.7 m/sec determined by Johnson for block 1 corresponds to 0.50-m 

radial distance using the seismic velocity of 2,640 m/sec and Equation 12. The crushing limit of 

4.5 m/sec that was determined using the factor for the Persson et al. [1994] crushing limit for 

block 3 corresponds to 0.50-m radial distance using the seismic velocity of 2,640 m/sec and 

Equation 12. 

Radial cracks in block 3 were counted per unit arc length at radial intervals from the 

blasthole (Figure 24). 

Figure 24. Block 3 photograph/quadrant overlay of postblast wire saw section near blasthole 

collar showing cracks. 

The number of cracks and arc angles are listed in Table 11. 

Table 11. Block 3 number of radial cracks visible in wire saw cut 

Distance 

(m) 

Number of 

cracks 

Aperture 

(degrees) 

Cracks per m 

arc length 

0.25 4 90 10.2 

0.50 4 90 5.1 

0.75 3 90 2.5 

1.00 2 55 2.1 

1.25 1 40 1.1 

32
 

There is a 

decrease in number of cracks per unit arc length with distance as shown in the plot in Figure 25. 
Figure 25. Block 3 postblast number of cracks vs. distance from blasthole. 

The number of cracks at the MAE Rd limit of 0.57 m using the regression equation in Figure 

25 is 4.9 cracks per meter of arc length. The PPV as measured using strain at the MAE Rd limit is 

3.3 m/sec. 

Discussion 

Figure 26 provides a summary of the measured damage, predicted damage, and the 

calculated MAE Rd practical damage limit for each block tested. The measured damage using the 

MVP and the crosshole seismic methods correlates with the MAE Rd practical damage limit. 

Results are dependent on the percentage cutoff of the damage curve asymptote for each dataset. 

The measured limit is based on 98% of the asymptote and is slightly higher than the MAE Rd. 

The Persson et al. [1994] and Johnson [2010] PPV limit assessments for crushing and seismic 

limits are less than the MAE Rd practical damage limit but are within 80% of the MAE Rd value. 

The number of radial cracks per meter of arc length estimated at the MAE Rd limit is 12.3, 

11.2, and 4.9 for blocks 1, 2, and 3, respectively. The lower crack number for block 3 may 

indicate differences in parameters including block strength, block aggregate size, block density, 

explosive emulsion, charge length, and charge stemming. Blocks 1 and 2 had similar parameters 

and comparable crack results. 

The measured PPV at the calculated MAE Rd practical damage limits for blocks 1, 2, and 3 

are 4.9 m/sec, 5.0 m/sec, and 3.3 m/sec using the regression equation in Figure 15 and the 

seismic velocities. 

The three block experiment results are data limited and preclude statistical analysis but 

provide a better understanding of the MAE Rd limit if used for blast design. When the MAE Rd 

limit is applied to a buffer hole design location, the MAE Rd limit would be expected to have 

radial cracks extending up to and beyond the limit, yet the primary damage expected by Persson 

et al. [1994] and by Johnson [2010] should be within the limit. The majority of damage as 

measured seismically using either the MVP or the crosshole techniques would occur up to the 

MAE Rd limit. 

33
 
 

Figure 26. Summary of damage limit results of block experiments compared to MAE Rd. 

Buffer Hole Design Concept 

The buffer hole concept is an adaptation of the practical damage limit for design. In the 

buffer hole design, the contour sector is determined by the buffer holes and not by the perimeter 

holes. The buffer concept was first published by Hustrulid and Johnson [2008]. The contour strip 

of rock is bounded on the inside by the buffer row of holes and on the outside by the perimeter 

(contour) row of holes. The detonation of the buffer row will produce damage in the contour 

sector. The idea is to design the buffer holes so that their associated damage radius extends to the 

desired drift perimeter. Figure 28 shows the case when the radius: 



contour row “burden”

Rd

34
 

(14) 
Figure 27. Buffer row hole locations, buffer hole practical damage circles, 

and perimeter holes at the cusp locations. 

If this is the case, after the firing of the buffer holes, the amount of fresh/undamaged rock is 

only the small cusp of rock remaining between damage circles. With this in mind, the breaking 

demand on the perimeter holes is substantially reduced. This is a prime reason why the use of 

high-strength detonating cord often functions quite well as a perimeter control explosive because 

the primary function has become one of smoothing rather than primary breaking. 

The key to the buffer row approach [Hustrulid and Johnson 2008] is the assignment of a 

“practical” radius of damage (Rd) to each blasthole/explosive combination being considered for 

use in the particular rock mass. The following steps are used after calculating the MAE Rd 

practical damage limit: 

Step 1: Design the buffer row starting with buffer circles tangent to the abutment corners.
 

Step 2: Add the perimeter holes at the abutment corners and at the cusp locations.
 

Step 3: Design the lifters.
 

Step 4: Add the cut.
 

Step 5: Add stope B and C holes as required providing good energy coverage or based on 

powder factor experience.
 

35
 
 

Buffer hole spacing is suggested between 1.3 Rd and 1.5 Rd to provide area coverage of 

between 92.4% and 89.6% respectively. A plot of a wide range of buffer spacing factors and 

associated area coverages are shown in Figure 28. 

Figure 28. Range of buffer hole spacing conditions showing percentage of area covered by circles 

formed using the Rd damage radius. 

A high percentage area of coverage is 

advantageous for reducing the effort required for the perimeter charge. The 1.3 Rd and 1.5 Rd 

range is untested. 

Perimeter Charging 

The buffer hole design procedure results in perimeter holes located at the cusps between 

buffer hole damage circles. Charging of the perimeter holes is decoupled to reduce the charge 

amount so as to crack the rock just sufficiently enough to trim the remaining rock. Sanden [1974] 

applied the force-equilibrium approach in developing a hole spacing (S) relationship for 

presplitting. The same perimeter spacing is suggested by Hustrulid and Johnson [2008] for 

contour blasting application using rock strength and explosion pressure using the equation 















 



t

tw

h

P

rS





2

(15) 

where S = perimeter hole spacing (m), 

rh = blasthole radius (m), 

Pw = explosion pressure exerted on the blasthole wall (MPa), and 

ζt = rock tensile strength (MPa). 

36
 
 

If tensile strength is not known, a factor of, for example, 1/12 the compressive strength could 

be used. According to Hustrulid and Johnson [2008], Pw is designed to be less than the rock 

compressive strength to prevent rock crushing. However, the spacing result will likely be 

different than the buffer hole concept design spacing. Alternatively, with the spacing already 

determined by the buffer row concept design, the correct explosive amount could be determined 

by rearranging the Sanden [1974] equation where 

t

h

t

w

r

S

P 





2

(16) 

The explosive and the calculated explosion wall pressure, using the pressure calculation 

described by Hustrulid and Johnson [2008], were chosen to match the pressure determined from 

Equation 16. Hustrulid and Johnson [2008] suggest the maximum wall pressure equal to the 

compressive strength of the rock to prevent crushing. 

Discussion 

In the buffer hole design concept the placement of the buffer holes determines the designed 

perimeter and the demand on the perimeter holes is substantially reduced. Perimeter charging 

options for the buffer design concept will require decoupling. Without perimeter hole 

decoupling, damage typical of an aggressive blast may result. Explosives manufacturers have 

products specifically suited for perimeter control; two examples are: (a) a detonation cord which 

comes in various charge concentrations and (b) emulsion-based trim cartridges or continuous 

charges. An alternative option is to use pumped emulsion or blown ANFO as a bottom charge. 

This reduces the overall energy in the perimeter hole. Another solution is to use ANFO fully 

coupled, in addition to tracing the ANFO with a detonation cord. The ANFO will not reach its 

velocity of detonation (VOD) potential, and the gas volume created from ANFO will split the 

rock similarly to a decoupled charge. 

The use of detonators, boosters, or cartridges for initiation of the column charge is not 

considered in the buffer design. The length of the blastholes and the resulting length of the blast 

round are not considered in the buffer design. The design considers the buffer and perimeter 

holes to be drilled parallel to each other. Rock structure is not considered in the buffer row 

concept design, even though the rock structure can affect the blast outcome. Singh and 

Narendrula [2007] have done considerable research on the influence of rock structure. Research 

indicates that blasthole spacing should be less than the joint spacing because radial cracks will 

likely arrest at a joint surface. Hustrulid [1999a] indicates that closer joint spacing requires more 

blastholes of smaller diameter. 

37
 
Analysis of Successful Perimeter Control Designs
 

Introduction 

Successful controlled blast designs provide as-built results that are equivalent to as-designed 

results. The perimeter is controlled and typically results in half-barrel remnants from the 

perimeter holes indicating no overbreak and only minimal fracturing into the perimeter. Five 

successful perimeter control designs were chosen to show the relevance of the buffer row 

concept and any differences between both design concepts. The following sections provide a 

comparison of the buffer hole design concept to the five successful designs including a 

comparison of buffer hole placement, perimeter hole spacing, and hole charging. 

Example of a Successful String-Loading Perimeter Control Design 

Decoupling through the process of string-loading bulk emulsion in the perimeter holes is 

thought to be a very appealing alternative which should be considered by U.S. mining 

companies. Although not available in the United States at the time of this writing, string loading 

is a common method for perimeter control elsewhere in the world. This technique offers a 

relatively simple means for applying perimeter control for use in production holes and is 

desirable for wet conditions, thereby reducing the number of types of explosives in the round to 

just one. The variable speed charging hose retraction rate offered by the string-loading 

equipment allows the miner to adjust the perimeter decoupling ratio. Figure 29 illustrates the 

string-loading process. 

Figure 29. Photograph and diagrams of string-loading bulk emulsion using constant pumping and 

variable rate extraction of the injection hose [Fauske 2003; Hustrulid 2008]. 

38
 
A successful controlled blast design using string loading is currently being applied at a mine 

in Australia (see Figure 30). This design will serve as the basis for analysis using the buffer row 

concept because the position of the buffer row is parallel to the perimeter and approximately at 

the correct damage distance from the perimeter. This is considered a successful design because 

the as-designed limit is equivalent to the as-drilled limit and the as-built limit. Damage beyond 

the as-built limit is subjective because no measure of radial cracks was made. Overbreak has 

been minimized and the perimeter is defined by the perimeter blasthole half-barrels. 

The basis used for the design, especially the buffer row distance from the perimeter, is 

uncertain. It may in fact be based on a standard empirical-based design or a special design from 

the mine’s expert consultant. This design was selected as the one that most closely follows the 

buffer row design. 

The authors of this report visited the Australian mine site and collected geotechnical and 

design data. 

39
 
Figure 30. Example showing drill hole locations of a successful blast design in Australia that uses 

string-loaded emulsion. 

Explosive Properties 

The round was drilled using a computer-controlled drill jumbo. The round was charged using 

site-sensitized emulsion. The explosive properties include energy at 3.1 MJ/kg (740 kcal/kg), gas 

3 volume at about 950 L/kg, density at 0.85 g/cm , VOD equal to 4,300 m/s, and relative weight 

strength (RWS) or sANFO equal to 0.84. The buffer holes were fully charged and the perimeter 

holes were string loaded to 50% of the hole cross-sectional area (0.034 m in diameter). The 

adiabatic constant γ is assumed to be 3.0. 

Drift Geometry 

The drift geometry is 5.5 m wide, 4.8 m to the abutment corners, and 6 m to the crown. Blasthole 

length was 6 m and the hole diameters were 48 mm. The buffer row Rd was 0.54 m. A buffer row 

is prominent with a spacing of 0.85 m (1.57 Rd). The perimeter row spacing was 0.65 m (0.2 Rd). 

40
 
Rock Properties 

3 The rock type is monzonite with a density of 2.8 g/cm and a Poisson’s ratio of 0.28. The rock 

friction angle is 45°. The compressive strength is 150 MPa and the tensile strength is 22 MPa 

from Brazilian tests. 

Blast Damage Calculations 

The calculated MAE Rd practical damage limit for the fully coupled buffer holes is 0.54 m or 

22.5 blasthole radius. 

Discussion 

Table 12 compares the calculations of the buffer design approach to the successful design. 

Being a successful design implies that there is no improvement needed.

Table 12. Buffer row design concept parameters and damage 

compared to successful Australian design 

Design Rd 

(m) 

Buffer spacing Perimeter spacing Sanden perimeter Pw 

(MPa) 

MAE 

Successful 

design 

0.54 

0.54 

1.3 Rd to 1.5 Rd 

1.57 Rd 

1.3 Rd to 1.5 Rd 

1.2 Rd 

300 to 349 

275 

(598 covolume 

wall pressure) 

 However, it is useful to 

compare the successful design to provide validation to the buffer design approach. The 

successful design shows several important relationships using the applied Rd practical damage 

radius. The successful design perimeter burden or buffer hole Rd is equal to the MAE calculated 

Rd of 0.54 m. 

The buffer hole spacing measured on the design is 0.85 m or 1.57 times the designed Rd of 

0.54 m and is slightly larger than the suggested buffer hole spacing range of 1.3Rd to 1.5Rd. The 

perimeter hole spacing is 0.65 m and is less than the buffer hole spacing of 0.85 m. 

Spacing of the perimeter holes at the cusps, which is part of the buffer row design concept, 

occurs for 9 of the 21 perimeter holes in the successful design. If the buffer design concept was 

applied (buffer holes alternate with perimeter holes), the result would be a larger spacing of 

perimeter holes and less perimeter holes than used in the successful design. 

The Sanden [1974] perimeter charge wall pressure equation results are less than the 

decoupled, string-loaded wall pressure of 598 MPa of the successful design [Hustrulid 2010]. 

The decoupled pressure calculation is shown in Table 20 in the Appendix. The Sanden [1974] 

calculated pressures needed for the various perimeter spacings all exceeded the compressive 

strength of 150 MPa. 

41
 
ch e ch e decoupled LrLr hole

2

arg

2

 arg  

Example of a Successful Bottom Charge Perimeter Control Design 

This successful development drift design is used at a block cave operation in the United 

States. 

Explosive Properties 

The blastholes were charged with ANFO. A detonator and booster were used for initiation. 

For perimeter control, ANFO was used for the back perimeter holes by charging the bottom third 

of the hole and leaving the upper two thirds of the hole empty. The collar was typically plugged 

to provide an air deck between the collar and the bottom charge. The result was a decrease in the 

overall explosion pressure over the uncharged length of the hole, even though the bottom third 

was fully coupled. The perimeter charge explosion pressure calculation requires an assumption 

that the bottom charge explosion pressure is reduced over the uncharged hole length. The fully 

coupled bottom charge volume is distributed along the entire hole length as: 

(17) 

where rcharge = bottom charge radius (m), 

Lcharge= bottom charge length (m), 

rdecoupled = assumed charge radius for decoupling calculation (m), and 

Lhole = assumed charge length for decoupling calculation = hole length (m). 

The equation is solved for rdecoupled and a decoupled explosion pressure is calculated (Appendix 

Table 21). In the case for the successful design bottom charge, with the average hole length of 

4.4 m and a bottom charged length of 1.5 m, the assumed decoupled charge diameter is 0.014 m. 

Drift Geometry 

The drift design (Figure 31) is nominally 4.28 m x 4.27 m in cross section with an arched 

roof and an abutment height of 2.1 m. A line of buffer holes parallels the perimeter holes 

defining a perimeter burden of 0.76 m. Blastholes were 0.048 m in diameter and 4.4 m long. The 

perimeter burden (buffer hole Rd) is 0.73 m and the buffer hole spacing is 0.8 m or 1.1 Rd. The 

perimeter hole spacing is 0.73 m (1.0 Rd). 

Rock Properties 

The mine geology consists of quartz vein stock works. The rock type was primarily 

3 porphyritic with a compressive strength of 77 MPa, a density of 2.5 kg/m and a Poisson’s ratio 

of 0.22. The rock mass quality was determined to have a Rock Mass Rating (RMR) of 48. 

Blast Damage Calculations 

The calculated MAE Rd practical damage limit for the fully coupled buffer holes is 0.68 m or 

28.3 blasthole radius. 

42
 
Figure 31. Example of a successful blast design in the United States with 

bottom-charged perimeter holes. 

Discussion 

Table 13 compares the buffer design approach calculations to the successful design. The 

successful mine design employs a common perimeter control practice of bottom charging the 

perimeter holes to reduce the explosion energy and in turn reducing perimeter damage. Though 

not as effective as a full-length decoupled charge, the bottom charging method does reduce 

damage compared to fully coupled charging and can be considered a successful blast design. 

Perimeter damage along the fully coupled section (bottom charge) of the perimeter holes is 

expected to have greater damage than the air-decked section. 

43
 
 

The successful design buffer hole locations compared well with the damage model. The 

successful perimeter hole spacing is less than what would be expected for the buffer design 

concept of 1.3 Rd to 1.5 Rd. However, the successful perimeter spacing of 1.0 Rd is nearly 

equivalent to the buffer spacing of 1.1 Rd for the successful design indicating perimeter holes are 

at the cusps. Equivalent buffer and perimeter spacing is a requirement for the buffer design 

concept. 

The successful design provides additional insight into possible adjustments for the buffer row 

design concept. It might be worthwhile to try a buffer hole spacing less than the conceptual range 

of 1.3 Rd to 1.5 Rd. The successful perimeter hole spacing would then align with the new buffer 

hole cusp locations. The actual decoupled pressure calculation of 56 MPa is less than the 

required calculated pressures by Sanden [1974]. The actual decoupled pressure calculation is 

shown in Table 21in the Appendix. 

Table 13. Buffer row design concept parameters and damage compared to 

successful U.S. design 

Design Rd 

(m) 

Buffer spacing Perimeter spacing Sanden perimeter Pw 

(MPa) 

MAE 

Successful 

design 

0.68 

0.73 

1.3 Rd to 1.5 Rd 

1.1 Rd 

1.3 Rd to 1.5 Rd 

1 Rd 

112 to 130 

91 

(56 covolume 

wall pressure) 

Example of a Successful Swedish Decoupled Cartridge Perimeter Control Design 

This design example is an underground zinc/silver operation. The mine uses cut and fill 

stoping techniques in steep-dipping narrow veins in rock that is primarily limestone [Norling and 

Nord 2006]. Information for this analysis is based on design information from Norling and Nord 

[2006] and personal communication from Marklund [2011]. 

The mine had previously used contract miners for drift driving. There was a concern about 

overbreak problems and the apparent poor rock quality after blasting. A new blast design was 

implemented using in-house resources for both drilling and blasting. The new blast design is 

shown in Figure 32. The design shows the use of perimeter control by using: (1) close-spaced 

perimeter holes and (2) a buffer row parallel to the perimeter row. A two-relief hole burn cut is 

used with stoping B holes on each side of the burn cut, stope C holes above the burn cut, knee 

holes below the burn cut, and lifters to define the floor. 

Explosive Properties 

The properties of the explosives used for this blast design are shown in Table 14. Note that, 

with the exception of ANFO, all of the charges are decoupled. 

44
 
 

The published relative weight strengths of the explosives are shown in Table 14. The sANFO 

values were also calculated from the gas volume and energy of the explosives as provided by the 

mine. Properties for the Emulit 22 gas volume were not available from the mine. The following 

Swedish equation is used to calculate the relative weight strength with respect to a base 

explosive: 

0 0 6

1

6

5

V

V

Q

Q

s 

(18) 

where	 s = Relative weight strength,
 

Q = Explosive energy (MJ/kg),
 

Q0 = Base explosive energy (MJ/kg),
 

3 V = Explosive gas volume (m /kg), and
 

3 V0 = Base explosive (ANFO) (m /kg).
 

In this example ANFO has been used as the base explosive. The referenced weight strengths 

compared well with the calculated values. 

Table 14. Successful Swedish controlled blast design explosive properties 

Hole Type Diameter 

(mm) 

Density sANFO, 

published 

weight 

strength 

VOD 

(km/sec) 

Q 

(MJ/kg) 

V 

(m3/kg) 

sANFO, 

calculated 

weight 

strength 

Perimeter Gurit 17 17 1,000 0.85 2.4 3.4 .930 0.87 

Buffer Emulit 22 22 1,130 nd* 5 2.4 1.12 0.69 

Stope and 

cut 

ANFO 48 850 1.0 2.2 4.0 .970 1.0 

Lifter Dynamex 

32 

32 1,450 1.13 4.5 4.5 .890 1.09 

Sources: Norling and Nord [2006], Persson et al. [1994], Holmberg [1992], and Hustrulid and 

Johnson [2008]. 

*nd = not determined 

Geometry 

The drift dimensions are 5 m wide, 4.9 m to the abutment corners, and 6 m in overall height. 

All holes are 4 m in length and 0.048 m in diameter. The perimeter burden (buffer hole Rd) is 0.5 

m, and the buffer hole spacing ranges from 0.7 m to 1.0 m. The perimeter hole spacing is 0.7 m. 

45
 
Rock Properties 

The rock type is limestone in generally good ground based on the RMR of 75 or higher. 

Table 15 lists some of the important rock properties [Markland 2011]. 

Figure 32. Example of the successful blast design in Sweden with decoupled perimeter cartridges. 

46
 
 

Table 15. Properties of rock at the example mine 

Property 

Rock type 

RMR 

Modulus of elasticity 

Poisson’s ratio

UCS 

Density 

Sonic velocity 

Friction angle

Value 

Limestone 

>75 

60 GPa 

0.12 

100 MPa 

2,700 kg/m3 

4,860 m/sec 

Unknown (45° assumed)

Blast Damage Calculations 

The calculated MAE Rd practical damage limit for the decoupled buffer holes is 0.30 m or 

12.5 blasthole radius. The MAE calculation presented in this report does not account for 

decoupling. The MAE result was multiplied by the decoupling ratio as described by Hustrulid 

[2010] to best estimate the decoupled MAE Rd value. 

Discussion 

The buffer design approach calculations are compared to the successful design in Table 16. 

The MAE damage calculation for Emulit 22 for the buffer row Rd was 0.30 m and less than the 

successful design Rd of 0.5 m. The successful design’s 0.7-m perimeter spacing is correct for the 

buffer row design concept range of 1.3 Rd to 1.5 Rd. The successful design’s back buffer hole 

spacing also fell within the 1.3 Rd to 1.5 Rd design concept range. The rib buffer spacing was 

greater than the design concept range. Perimeter charge wall pressure was calculated to be 33 

MPa. The Sanden equation results were higher. The decoupled pressure calculation is shown in 

the Appendix in Table 22. 

Table 16. Buffer row design concept parameters and damage compared to 

successful Swedish design 

Design Rd 

(m) 

Rd/rh Buffer spacing Perimeter spacing Sanden Perimeter 

Pw (MPa) 

MAE 

Successful 

design 

0.3 

0.50 

12.5 

22.2 

1.3 Rd to 1.5 Rd 

1.4 Rd to 2.0 Rd 

1.3 Rd to 1.5 Rd 

1.4 Rd 

59 to 70 

113 (33 covolume 

wall pressure) 

47
 
Example of a Successful Canadian Controlled Blast Design 

The successful Canadian design example is from an underground nickel mine using cut-andfill and vertical block mining methods [Marshall et al. 1983]. This successful development 

controlled blast design has been described by Sutherland [1989] and Cudmore [2001]. The 

design utilized a buffer row approach with careful perimeter blasting using an Orica Powersplit 

perimeter control explosive. The perimeter and buffer hole placements are shown in Figure 33. 

Figure 33. Example of a successful blast design in Canada with decoupled perimeter cartridges 

(shown are only the locations of the buffer, perimeter, and lifter holes). 

48
 
Explosive Properties 

The perimeter control explosive that proved most effective for the successful Canadian 

design was the 22-mm-diameter, continuous-packaged Orica Powersplit slurry product. The 

buffer and production holes were charged with AmexTM ANFO. The explosive properties are 

listed in Table 17. 

Table 17. Explosives used for Canadian successful design 

Explosive 

Density 

(kg/m3) 

Explosive 

diameter 

(m) 

Published 

VOD 

(km/sec) RWS 

Orica Amex 

Orica Senatel Powersplit 

(traced with 10 g/m det cord) 

1,000 

1,300 

0.048 

0.022 

3.3 

7 

1.1 

1.09 

Geometry 

According to Sutherland [1989], the perimeter hole burden and spacing were based on the 

hole diameter where k = 15 to 16 and the burden was 0.8 of the spacing as described in the 

Modified Ash Energy (MAE) Approach section. The development drift dimensions were 4.3 m 

wide, 3.8 m to the abutment height, and 4.4 m in total height. The blastholes were 45 mm in 

diameter and 3.6 m in length. The perimeter burden (buffer hole Rd) was 0.66 m. The buffer hole 

spacing was 0.95 m for the rib buffer holes and 0.71 m for the back buffer holes. Perimeter hole 

spacing was 0.8 for the ribs and 0.76 m for the back. 

Rock Properties 

The rock was schistose gneiss with assumed rock properties of 150 MPa compressive 

strength, 0.25 Poisson’s ratio, and a 45° friction angle. 

Blast Damage Calculations 

The calculated MAE Rd practical damage limit for the fully coupled buffer holes is 0.62 m or 

27.6 blasthole radius. 

Discussion 

The buffer design approach calculations are compared to the successful design in Table 18. 

The successful buffer row Rd was comparable to the MAE results. The buffer spacing of 1.1 Rd 

to 1.4 Rd was also comparable to the suggested buffer design spacing of 1.3 Rd to 1.5Rd. 

Perimeter spacing was slightly less than the suggested buffer design spacing. The Sanden [1974] 

perimeter explosion pressure for the successful design was comparable to the Sanden calculated 

pressures for the MAE buffer design approach calculations. The actual decoupled perimeter hole 

wall pressure was calculated to be 612 MPa. This pressure seems high and is the result of a 

higher VOD and density values for the traced detonation cord combined with the Senatel 

emulsion explosive. The decoupled pressure calculation is shown in the Appendix in Table 23. 

49
 
Table 18. Buffer row design concept parameters and damage compared to 

successful Canadian design 

Design Rd 

(m) 

Buffer spacing Perimeter spacing Sanden Perimeter Pw 

(MPa) 

MAE 

Successful 

design 

0.62 

0.66 

1.3 Rd to 1.5 Rd 

1.1 Rd to 1.4Rd 

1.3 Rd to 1.5 Rd 

1.2 Rd 

211 to 245 

208 (612 covolume 

wall pressure) 

Example of a Successful Spanish Railway Tunnel Construction Controlled Blast Design 

This railway tunnel design is the initial upper cut of the tunnel. The site was visited by the 

authors of this paper in 2009, and geotechnical and design data were collected. As seen in the 

upper cut design in Figure 34, the buffer holes are at a constant distance from the perimeter and 

are equally spaced. 

Figure 34. Example of the successful railway tunnel blast design in Spain. 

Explosive Properties 

The perimeter holes were bottom charged with two sticks of a cartridge explosive and a 100-

gram/m detonation cord with an assumed VOD of 7 km/sec. For the explosion pressure 

calculation, the diameter of the contained PETN in the detonation cord was determined to be 

3 0.0095 m assuming a density of 1,400 kg/m . The production and buffer holes were charged with 

a gelatin nitroglycerin-based explosive, Goma 2 E-C, widely used for industrial applications in 

Spain. The bottom 1.2 m of the holes was charged with 40-mm-diameter cartridges, and the 

remaining 0.8-m charged length was loaded with 32-mm cartridges. The similar properties of the 

Orica Powerfrac were used for the buffer damage model calculations. 

50
 
 

Geometry 

The tunnel heading was 13 m wide and 5.34 m high. The blastholes were 3 m long and 45 

mm in diameter. The perimeter burden (buffer row Rd) was 0.57 m. The buffer row spacing was 

1.17 m. The perimeter row spacing was 0.62 m. 

Rock Properties 

The rock type was limestone with RMR values ranging from 35 to 60. The compressive 

strength ranged from 25 MPa to 60 MPa. The friction angle ranged from 28° to 39°. Poisson’s 

3 ratio was assumed to be 0.25. Rock density was assumed to be 2.65 g/cm . 

Blast Damage Calculations 

The calculated MAE Rd practical damage limit for the fully coupled buffer holes was 0.67 m 

or 29.8 blasthole radius. 

Discussion 

The buffer design approach calculations are compared to the successful design in Table 19. 

The successful tunnel design provides additional insight for comparison to the buffer row design 

approach. The design had a distinct buffer row with a consistent distance Rd of 0.57 m to the 

perimeter. The MAE calculated Rd was greater than the successful design Rd. 

The successful buffer hole spacing of 2.1 Rd was higher than the concept design suggested of 

1.3 Rd to 1.5 Rd. 

The successful design required closer perimeter spacing than suggested by the concept 

design. The calculated Sanden [1974] perimeter wall pressure Pw for splitting was 65 MPa and 

less than the actual wall pressure of 84 MPa. The MAE calculated splitting pressure was 

reasonably close to the pressure calculated for the successful design perimeter charge. The 

decoupled pressure calculation is shown in the Appendix in Table 24. 

Table 19. Buffer row design concept parameters and damage compared to 

successful tunnel design 

Design Rd 

(m) 

Buffer spacing Perimeter spacing Sanden perimeter Pw 

(MPa) 

MAE 

Successful 

design 

0.67 

0.57 

1.3 Rd to 1.5 Rd 

2.1 Rd 

1.3 Rd to 1.5 Rd 

1.1 Rd 

91 to 107 

65 (84 covolume 

wall pressure) 

51
 
Discussion 

The average buffer row Rd ratio (Rd/rh) for the successful designs was 24.4. This is only 

slightly less than the MAE Rd approach base value of 25. In most of the comparisons the damage 

model Rd values correlate with the successful design Rd values. The buffer hole Rd values that 

were appreciably different were the successful Swedish and Spanish designs. Discrepancies in 

the buffer row concept design include the slight differences in the buffer hole spacing and 

consistently fewer perimeter holes. The Sanden calculations for perimeter hole pressure are 

interesting in that two of the successful designs had much higher pressures than required for 

splitting between perimeter holes. Conversely, three of the successful designs had lower 

pressures than suggested by Sanden [1974], although perimeter smoothing was successful. The 

successful low perimeter hole wall pressures indicated that the requirement of splitting between 

perimeter holes is unnecessary and only cusp removal or local perimeter smoothing was 

required. 

Design Guidance 

Some guidance is suggested for the user applying the buffer row design approach. These 

guidance points are based on the data and analysis presented in this report. Guidance points 

relating to the buffer hole Rd damage radius and spacing are: 

	 The calculated Rd is a starting point for design. The calculated result should be close 

to what will provide the best damage control result but should not be considered the 

absolute result. 

	 Trial blasts are suggested to optimize the controlled blast design. 

	 The MAE Rd damage model is a sound empirical approach for average rock 

conditions; Rd calculations from explosives used in the successful designs had in all 

cases, except in one, good correlation with actual Rd distances. 

	 Any Rd/rh user results that are outside the range of 20 to 30 for buffer holes should be 

considered suspect. The range of Rd/rh values used in the successful designs was 20.8 

to 30.4. 

	 Buffer hole spacing is suggested within a range of 1.3 Rd to 1.5 Rd. This is a valid 

range because the successful designs vary from 1.1 Rd to 2.1 Rd. The lower end of the 

suggested range is recommended to provide closer spacing of the perimeter holes 

which are located at the cusps between buffer holes. 

Guidance points relating to use of the proper perimeter hole charging and spacing are: 

	 The perimeter hole spacing is based on the buffer holes spacing. 

	 Charging of the perimeter based on Sanden [1974] will assure the user that a high 

enough hole pressure is used to split the perimeter in the event the buffer row did not 

damage the perimeter burden sufficiently. 

	 Perimeter hole spacing as described in the buffer row design method is suggested 

within a range of 1.3 Rd to 1.5 Rd which is the same as the buffer hole spacing. The 

successful designs’ perimeter spacing varied from 1.0 Rd to1.4 Rd indicating that the 

suggested spacing should be closer to 1.3 Rd or even lower. In support for closer than 

52
 
1.3 Rd spacing for the perimeter row is the point that the lowest buffer spacing in the 

successful designs was 1.1 Rd, which would result in equally close spacing for the 

perimeter holes if the buffer row concept was used. Further, the percentage of area 

covered by the buffer hole MAE Rd circles would be greater as the circle spacing 

decreases. A compromise in hole spacing to assure a successful controlled blast 

design that would accommodate both the buffer spacing and the perimeter spacing 

would be a narrow range around 1.3 Rd. 

Conclusions 

Overbreak is a concern in drift operations because of the associated loose rock, additional 

scaling requirements, and the increase in potential for fall-of-ground accidents. Excessive 

overbreak is a phenomenon that occurs because perimeter control is not generally applied. 

Today, available drilling and blasting technology has developed to the point that there is no 

reason why the rock mass cannot be “cut as with a knife”, if so desired [Kvapil 2008]. A 

practical damage estimator Rd using the buffer row design method was presented as a solution to 

control the perimeter and reduce scaling requirements. 

Three large block experiments confirmed the expected radial damage from fully coupled 

charges with measurements made both visually and with acoustic measurements. These tests 

confirm a close relationship between the practical damage radius Rd and the measured damage 

limits when applying the MAE approach. 

The step-by-step algorithm for placement of buffer and perimeter holes was briefly described 

utilizing the Rd practical damage limit. 

The practical damage estimator, Rd, was applied to actual successful perimeter control 

designs to affirm the use of the estimator for design. The estimator was shown to be valid in each 

example. It was found that the spacing of the buffer holes and the perimeter holes varies in the 

successful designs. A buffer design with buffer hole and perimeter hole spacings of 1.3 Rd is a 

good starting point and based on the successful designs. 

The buffer row design concept provides a simple, and technically sound, method for 

assigning the damage radius for a particular explosive, hole diameter, and rock density 

combination. 

Recommendations 

Controlled blasting is recommended over aggressive blasting methods to reduce perimeter 

damage, scaling, and fall-of-ground accidents. The design, whether it is the buffer row concept 

described in this report or another design approach, must be engineered and implemented with 

the “buy-in” of the miner. Precision drilling is necessary to achieve as-drilled performance that is 

equal to as-designed performance. The jack leg drill cannot easily achieve this performance 

result; however, drill jumbos with precision control that can achieve as-designed results are 

recommended. 

53
 
Deep mine stress and existing rock structure could affect the extent of blast damage. For 

example, the length of radial cracks could terminate at existing rock structure preventing further 

damage. High-stress conditions could alter the circular radial cracking pattern. Research into 

these areas is recommended. 

Acknowledgments 

The National Institute for Occupational Safety and Health (NIOSH) engineers who 

contributed to the controlled blasting research during the field studies and blast damage model 

development include Jami Dwyer, Douglas Tesarik, Joel Warneke, Edward McHugh, Chuck 

Kerkering, Yasser Akbarzedah, Rimas Pakalnis, Tom Brady, and Mark Kuchta. Publication 

outputs from the controlled blasting for safety project are listed in the Bibliography section 

following the References section. Other NIOSH technical staff who contributed to this research 

project includes Steve Ward, Ron Jackshaw, Mike Stepan, Paul Pierce, and Richard Rains. A 

special acknowledgment is given to the mines and their staff who provided information. 

54
 
References 

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Blasting Technique, International Society of Explosives Engineers, January 27–30, 2008. New 

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damage zone in drifting as demonstrated using concrete block data. Blasting and Fragmentation 

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59
 
60
 
Appendix A:
 

Perimeter Hole Covolume Wall Pressure Calculation Method
 

Used for Successful Design Examples 

61
 
 Pv  nRT 

  PP  eehw

 

 

Perimeter Hole Covolume Wall Pressure Calculation Method
 

Used for Successful Design Examples 

The pressure on the borehole wall for decoupled charges is explained in Hustrulid and 

Johnson [2008]. Generally, the explosion pressures are much higher than the compressive 

strength of the rock being blasted. Although this is desired when fracturing the rock in the 

interior part of the drift round, it is not true for the perimeter holes when perimeter control 

blasting is to be used. 

The first design requirement for these holes is to keep the borehole wall pressure less than or 

equal to the unconfined compressive strength. This is normally accomplished by using decoupled 

charges. The explosion pressure applies at the outer boundary of the charge. To reach the 

borehole wall, the explosive gases must expand. 

For ideal gases (gases at atmospheric pressure and room temperature), the standard 

expression relating pressure, volume, and temperature is 

 (19) 

where P = pressure, 

v = specific volume, 

n = number of moles of gas present, 

T = temperature, and 

R = the Universal Gas Constant. 

Assuming isothermal expansion, one writes 

 (20)

where Pe = explosion pressure, 

ve = specific volume of the explosive, 

Pw = wall pressure, and 

vh = specific volume of explosive gases filling the hole. 

Assume that 

ρe = 0.82 g/cm3 

62
 
.0/1/1 82 .1 22 cm /g3

ve e 

.1 22 .3 95 L/kg

30

54

2 2

 











 

















e

e

h

h

d

d

  

 

 )(  nRTvP 

The specific volume of the explosive would be 

 (21)
 

For the case when dh = hole diameter = 54 mm and 

de = explosive diameter = 30 mm, 

the specific volume of the gases filling the hole is given by 

 (22)

Applying Equation 20 and assuming an explosion pressure of 1560 MPa, the wall pressure 

would be 

482 MPa

.3 95

.1 22 ,1 560  





























h

e PP ew





 (23) 

However, one cannot apply this approach for the very high-pressure, high-temperature 

explosive gases involved here. The relationship relating pressure, volume, and temperature is: 

 (24)

where P = pressure (atm), 

v = specific volume (L/kg), 

 = covolume (L/kg), 

n = moles/kg, 

R = universal gas constant = 0.08207 L-atm/(K-mol), and 

T = temperature (K). 

63
 
 



.0 473/

1.1



 e h

 

 )()( P P   eeehhw 

 h

e h





.0 473/

1.1



 

 e

e e





.0 473/

1.1



 

 

The expression 

 (25)

will be used to relate the covolume and the specific volume. Assuming as before that the 

expansion of the gases in the borehole occurs isothermally, one can write 

 (26)

where 

 (27)

 (28) 

Substituting the appropriate values, one finds that 

1.11.1 .0 976 .0 473/ .0 473 .3/ 95  

 

e e

 h h

1.11.1 .0 718 .0 473/ .0 473 .1/ 22  

 

e e

 e e

The wall pressure with the covolume correction becomes 

263 MPa

.3 95 .2 974

.1 22 .0 502 ,1 560

)(

)(

 





































hh

ee PP ew





As can be seen, the covolume correction has a major effect. If the compressive strength is, 

for example, 

c = 200 MPa 

64
 

one would expect to see crushing around the hole. 
Table 20. Covolume spreadsheet calculation for Australian design 

Covolume 

calculation 

Value Units Notes Formula 

dh 0.048 m Blasthole diameter None 

de 0.034 m Explosive diameter None 

Evod 4.3 km/sec Velocity of detonation 

for the explosive 

None 

Edensity 850 kg/m3 Density of the explosive None 

v_e 1.176471 -- Specific volume of 

explosive 

1/(Edensity/1,000) 

Ve 0.440629 -- Volume term v_e-1.1*EXP(−0.473/v_e) 

vh_ 2.3448 -- Explosive gas specific 

volume 

v_e*(dh/de)^2 

Vh 1.445746 -- Volume term vh_-1.1*EXP(−0.473/vh_) 

Pewall 598 MPa Wall pressure from 

decoupled charge 

1/8*Edensity*Evod^2*(Ve/Vh) 

Coupled 

pressure 

1,965 MPa Wall pressure if fully 

coupled 

1/8*Edensity*Evod^2 

Table 21. Covolume spreadsheet calculation for bottom charge design 

Covolume 

calculation 

Value Units Notes Formula 

dh 0.048 m Blasthole diameter None 

de 0.014 m Explosive diameter None 

Evod 3.9 km/sec Velocity of detonation 

for the explosive 

None 

Edensity 950 kg/m3 Density of the explosive None 

v_e 1.052632 -- Specific volume of 

explosive 

1/(Edensity/1,000) 

Ve 0.350785 -- Volume term v_e-1.1*EXP(−0.473/v_e) 

vh_ 12.37379 -- Explosive gas specific 

volume 

v_e*(dh/de)^2 

Vh 11.31505 -- Volume term vh_-1.1*EXP(−0.473/vh_) 

Pewall 56 MPa Wall pressure from 

decoupled charge 

1/8*Edensity*Evod^2*(Ve/Vh) 

Coupled 

pressure 

1,806 MPa Wall pressure if fully 

coupled 

1/8*Edensity*Evod^2 

65
 
 

 

Table 22. Covolume spreadsheet calculation for Swedish decoupled design 

Covolume 

calculation 

Value Units Notes Formula 

dh 0.048 m Blasthole diameter None 

de 0.017 m Explosive diameter None 

Evod 2.4 km/sec Velocity of detonation 

for the explosive 

None 

Edensity 1,000 kg/m3 Density of the explosive None 

v_e 1 -- Specific volume of 

explosive 

1/(Edensity/1,000) 

Ve 0.314557 -- Volume term v_e-1.1*EXP(−0.473/v_e) 

vh_ 7.972318 -- Explosive gas specific 

volume 

v_e*(dh/de)^2 

Vh 6.935683 -- Volume term vh_-1.1*EXP(−0.473/vh_) 

Pewall 33 MPa Wall pressure from 

decoupled charge 

1/8*Edensity*Evod^2*(Ve/Vh) 

Coupled 

pressure 

720 MPa Wall pressure if fully 

coupled 

1/8*Edensity*Evod^2 

Table 23. Covolume spreadsheet calculation for Canadian design 

Covolume 

calculation 

Value Units Notes Formula 

dh 0.045 m Blasthole diameter None 

de 0.022 m Explosive diameter None 

Evod 7 km/sec Velocity of detonation 

for the explosive 

None 

Edensity 1,300 kg/m3 Density of the explosive None 

v_e 0.769231 -- Specific volume of 

explosive 

1/(Edensity/1,000) 

Ve 0.174466 -- Volume term v_e-1.1*EXP(−0.473/v_e) 

vh_ 3.218373 -- Explosive gas specific 

volume 

v_e*(dh/de)^2 

Vh 2.268719 -- Volume term vh_-1.1*EXP(−0.473/vh_) 

Pewall 612 MPa Wall pressure from 

decoupled charge 

1/8*Edensity*Evod^2*(Ve/Vh) 

Coupled 

pressure 

7,963 MPa Wall pressure if fully 

coupled 

1/8*Edensity*Evod^2 

66
 
Table 24. Covolume spreadsheet calculation for Spanish design 

Covolume 

calculation 

Value Units Notes Formula 

dh 0.045 m Blasthole diameter None 

de 0.0095 m Explosive diameter None 

Evod 7 km/sec Velocity of detonation 

for the explosive 

None 

Edensity 1,400 kg/m3 Density of the explosive None 

v_e 0.714286 -- Specific volume of 

explosive 

1/(Edensity/1,000) 

Ve 0.146999 -- Volume term v_e-1.1*EXP(−0.473/v_e) 

vh_ 16.02691 -- Explosive gas specific 

volume 

v_e*(dh/de)^2 

Vh 14.9589 -- Volume term vh_-1.1*EXP(−0.473/vh_) 

Pewall 84 MPa Wall pressure from 

decoupled charge 

1/8*Edensity*Evod^2*(Ve/Vh) 

Coupled 

pressure 

8,575 MPa Wall pressure if fully 

coupled 

1/8*Edensity*Evod^2 

67
 
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