Explosives Today - Series 4, No 5
% of maximum potential movement
10 Delay (ms/m)
Figure 8 Conceptual effect of delay period on ground movement for Intra-row and Inter-row delay
Given any combination with the above factors, the blast layout has a critical effect on the success of the cast blasting. We now look at those parameters which can be varied with relative ease and which therefore provide the most convenient means for optimising results. The key principles for maximising throw are: • Ensuring freedom of movement of the face, • Providing adequate energy • Ensuring the holes in each row fire as close to simultaneously as possible • Providing sufficient timing between these rows to achieve proper relief of burden These can be implemented by attention to: • Blast Layout o Block staking (Hole coordinates) • Explosive energy • Back filling of holes • Timing between front holes • Timing between rows of holes • Angle of holes to bench • Stemming • Priming position
Blast layout The drilling pattern is described by the burden and spacing of the blastholes which, together with the explosive density and the hole diameter, determine the powder factor of the blast. While it is usually possible to increase the percentage of rock cast without changing powder factor at all, conversion to optimised cast blasting normally involves an increase of between 15 and 50% in explosives consumption. – Drilled burden For a given rock, hole diameter and explosive, face velocity is primarily determined by the burden, that is the distance from the blasthole to a free, unrestricted face. Figure 6, generated using SABREX shows a typical Velocity – Burden curve for 251mm holes in a 28GPa medium strength unjointed Sandstone, using oxygen-balanced ANFO at a density of 0.8 and maintaining a constant powder factor. Because face velocity is directly influenced by powder factor, the drilling spacing has been varied in this graph to maintain a constant powder factor for the range of burdens shown. The graph shows
clearly that smaller burdens result in higher face velocities even at the same powder factor. The first row of holes is where the success of a cast blast is normally determined: if movement here is restricted it holds back the entire bench. The normal problem encountered is that of maintaining the design burden at the toes of these holes. Pre-splitting is the only technique by which to achieve consistent burdens in the front row, but even with this it can be difficult to ensure that correct burdens are maintained (Figure 7). Backbreak at the crest misleads drillers into collaring the holes too far back, while accumulations of overburden at the bottom, which look minimal from a distance, combine to severely hamper the development of full velocity at the toe. A degree of paranoia, both about the collaring of the first row of holes and the quality of the entire split face, is therefore well justified. It is also essential to ensure that there is no loose muck lying against the face, as this has a severe damping effect on movement. On the other hand, inadequate burdens in the front row result in blow outs at the face, releasing pressure before much
Explosives Today - Series 4, No 5
Point of initiation
8 meters 3 meters Pre-split line – fired with main blast Figure 9 Example of a cast blast layout
kinetic energy (acceleration) can be imparted to the bulk of the rock. The “ideal” muckpile profile varies from mine to mine and can be approximated by varying the drilling pattern across the bench. In order to ensure good initial movement, the first three rows of holes may be given reduced burdens (of say 6m), while the following rows may have expanded burdens (say 8m) to retard movement and raise the level of the muckpile at the back of the blast. This will facilitate movement of the dragline onto the muckpile: there is no point in moving the overburden forward if it has to be mechanically pushed back again later. The final row of blastholes breaks back to the pre-split as well as forward to the previous row. The standoff from the presplit is usually about one third of the forward burden: too little damages the pre-split wall while too much will leave unbroken accumulations on the wall. – Drilled spacing Hole spacing should be sufficiently wide to minimise the powder factor but close enough to preserve fragmentation and promote combined forward thrust between the holes in each row.
Provided a staggered drilling pattern is adopted, it is not necessary to keep the Spacing:Burden ratio of 1:1.15 as advocated for optimum fragmentation. A number of factors need to be considered in choosing the hole spacing, but a range between 1.15 to 2.7 can be considered quite legitimate. If excessive spacing is adopted, the holes firing within a row will not work to together and will leave areas of ‘dead’ ground between them. This increases the task of following rows of holes. The maximum spacing to avoid this effect is dependent on the rock type and delay system in use, but lies between 40 and 60 hole diameters. It is normal to standardise the hole spacing. Thus for a layout with burdens increasing from say, 6m at the front to 8m at the back, a spacing of 10m might be maintained. In the back row, this would be reduced to 5m, with a reduced charge to ensure full breaking against the pre-split. This is something to be considered together with the regional blasting engineer. – Explosive energy Clearly, the greater the energy in the blasthole, the greater the movement
Line of lightly charged holes next to pre-split
of the burden. Throw energy can be increased by: • Raising the density of the explosive • Using an explosive with a higher weight strength • Maximising ‘gas’ energy as opposed to shock’ energy given out by the explosive
Timing It is not always appreciated that there is a vast difference in the effects of INTRAROW as opposed to INTER-ROW timing. INTRA-ROW timing refers to the placing of delays between each of the holes in a row parallel to the face: ground movement is far more positive when these delays are very short. On the other hand, the movement of back rows is blocked by short INTERROW delays. Figure 8 is a qualitative illustration of these effects. Closely controlled blasting trials, using accurate delay detonators, have clearly demonstrated that maximum throw takes place when all the holes in a row detonate at the same instant. This is quite at odds with the fairly common practice of placing the longest delays
Explosives Today - Series 4, No 5
along the front row and firing back holes instantaneously with these, in echelon. It is no wonder that attempts to increase throw by increasing these delays have shown no significant effect on the percentage thrown. It is often not acceptable to, due to ground vibration and airblast effects, to fire an entire row of fully charged blastholes on the same delay, so the simple solution of firing row by row can seldom be adopted. However, it has been shown that minimal reductions in the face velocity result if intra-row delay is not more than 3ms per metre of hole spacing, for fairly tough rock. If back rows are to achieve effective movement, sufficient time is required before they detonate both for the backward acting forces of the previous row of blastholes to diminish and the previous burden to begin moving. The required time lag is dependent both on the explosive loading and the rock type, but the longer the delay, the greater the possibility of cut-offs and therefore the more critical the type of initiation system used. International operations use as much as 200ms between rows spaced 6 to 8 m apart with in-hole delays, but 80 to 100ms is probably the maximum where detonating cord downlines are in use. Thus a cast blast laid on a 6 – 8m by 10m pattern should ideally be timed using up to 20ms along the front and about 80ms between rows. (Figure 9 illustrates this principle using Shocktube Clusters). The desired timing pattern is not always easily achieved with conventional chevron timing using detonating cord, but there are a number of alternatives which should be discussed with your regional explosives engineer.
Stemming Patently, the longer the explosives gasses are confined, the more momentum is transferred to the rock
mass. The early release of gases in blow-outs can be minimised by: • Ensuring stemming length is at least 12 hole diameters and preferably 20 hole diameters • If short collar lengths are necessary for fragmentation then either a low energy detonating cord, electric detonators or shock tube or EDDs should be used to avoid disrupting the stemming • Using a medium (1/6 to 1/12 hole diameter) crusher aggregate in at least part of the stemming column to improve its retention properties. While this is good advice, it is seldom practical due to logistical problems
Priming Position Although this has not been thoroughly researched, there is evidence that bottom priming increases the movement of the overburden, probably due to both: a) The momentum of the detonation process imparting an upwards velocity to the rock mass b) The explosives being retained for a longer period before venting
Conclusion There is much that can be done to optimise the position of broken ground when blasting. However tight control over drilling and loading and close monitoring of the results achieved is vital when doing comparative trials to determine a blasting procedure. AEL Mining Services Explosive Engineers based at the regional offices are available to help and advise on any problems related to cast blasting. This document replaces all previous Explosives Today on this subject including Series 3. No 7: March 1990 and Series 3. No 8: June 1990 (CVB Cunningham)
AEL Mining Services Limited (PTY) Ltd 1 Platinum Drive, Longmeadow Business Estate North Modderfontein, 1645 Tel: +27 11 606 0000 www.aelminingservices.com
Disclaimer: Any advice and/or recommendations given by AEL Mining Services Limited (“AEL”) in this publication, is given by AEL in good faith in order to provide assistance to the reader. AEL does not however: 1.1 warrant the correctness of its advice and/or recommendations; 1.2 warrant that particular results or effects will be achieved if AEL’s advice and/or recommendations are implemented; 1.3 accept liability for any losses or damages that may be suffered, as a result of a party acting, or failing to act, on the advice and/or recommendations given by AEL;1.4 accept liability for any acts or omissions of its employees. representatives and/or agents, whether negligent or otherwise. Copyright: All copyright that subsists in this publication together with any and all diagrams and annexures contained herein, which shall include all and/or any ideas, plans, models and/or intellectual property contained in this document vests in AEL. Any unauthorised reproduction, adaptation, alteration, translation, publication, distribution or dissemination (including, but not limited to, broadcasting and causing the work to be transmitted in a diffusion service) of the whole or any part of this document in any manner, form or medium (including, but not limited to, electronic, oral, aural, visual and tactile media) whatsoever, will constitute an act of copyright infringement in terms of the Copyright Act No.98 of 1978 and will render the transgressor liable to civil action and may in certain circumstances render the transgressor liable to criminal prosecution. This document remains the intellectual property of AEL. Intellectual Property: All ideas, concepts, know-how and designs forming part of this publication belong to AEL, save for where it is clearly indicated to the contrary.
Explosives Today Series 4 I No 5
Cast Blasting Ken Meiring, Senior Explosives Engineer
1. The fixed parameters Introduction
When laying out blasting operations, the basic requirements are that the rock should be broken to full depth, that the fragmentation should be within acceptable limits and that safe conditions should prevail. Increasingly, the need to maximise the utilisation of capital equipment is calling for an additional criterion: â€œThe broken ground should fall in a position and shape that optimises overall productivity and mining costs.â€? While some operations require the minimal possible movement of the ground, cast blasting has the opposite aim. The prime application in opencast strip mines is to blast the material to the mined out space and then use draglines to move the remaining material to expose the top of the coal seam to be
Figure 1 Cast Blasting to improve dragline productivity
Explosives Today - Series 4, No 5
20 45 GPa Rock
28 GPa Rock
15 GPa Rock
Face Velocity (m/s)
9 x 10
Figure 2 Influence of rock elasticity on face velocity
extracted. (Figure 1.) The further the blasted material moves in the blast, the less is the dragline effort to expose the coal seam and the faster the coal can be extracted from the pit. The benefits of cast blasting are thus: • Reduced costs, higher production rates o Improved Total cubic metre rate (Tcm) o Bulk cubic metre rate (Bcm) o Improved bucket factor and digging rates • Less wear and tear • Less machine rehandle • Minimal dozer gain • More blast gain created • Competent dragline pads • Well fragmented material to maintain high level angle of repose to prevent spoil fluffing (fallback) • More coal recovery On higher benches it is advisable to strip enough top soil prior to drilling to avoid spoil constraint to be created at the end cuts and ramps. There are many factors influencing the success of attempts to achieve high throw factors and these can be classified into those which are
8 x 11
7 x 13
6 x 15
Drilling Pattern (m)
relatively hard or impossible to alter and those which can be varied with relative ease from blast to blast. Much of the information is gleaned from personal experience and the use of AEL’s computer blast simulation programs, which can be used to test the effect of various parameters on face velocity and muckpile shape. Because the fixed parameters are either not controllable or, once established, involve major disruption for change they are of critical importance. They should be carefully considered at the planning stage: • Geological Wireline information • Coal seam thickness and dip • Bench layout o Bench height o Bench width o Blasthole diameter o Drilling angle • Utilization and availability • Spoil constraints • Ramp blasting • Rehandle percentage • Dozer assist/Dozer gain • Maintenance schedule to be incorporated in the planning • Dragline/machine walkways
• Environmental constraints i.e. Mining closer to pan, boundary lines, water accumulation etc. • Inclusion of stab holes in critical areas e.g. dykes, fault areas, etc.
Rock characteristics The three most significant geological factors are the Vertical Joint Spacing (for uncemented joints), the Young’s Modulus of the intact rock and Rock Density, in that order. A rough guide to the way in which these influence the conversion of explosives energy into face velocity is given under each heading. – Vertical joint spacing Range
Diminishment of throw
<0.5 x Burden
0.5 – 1.5 x Burden
>1.5 x Burden
Open jointing which intersects the direction of movement has the effect of absorbing the initial movement of the rock and dissipating its kinetic energy in frictional processes. This effect is obviously increased if joint spacing is
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HIGH BENCHES Ground in partial tension Greater trajectory
LOW BENCHES Less of bench height charged Greater effect of toe drag Reduced trajectory
Figure 3 High Benches throw better
closer relative to the drilling pattern. Near horizontal jointing is less critical, unless the jointing is such that it provides early release of the explosives gases from the confinement of the blasthole: this can be detected by high speed photography showing premature venting from the face at prominent geological structural features. In addition to the above, general rock competence and uniformity is important in as much as it is vital to maintain an even burden on the front row of holes. An incompetent rock face is prone to slumping and slabbing, which inevitably means that the collaring positions of the front holes are too close to the edge of the bench for safe drilling. As a result these are set back, increasing the burden at the toe. This greatly reduces the likelihood of adequate throw: alternatively, if the holes are collared to achieve the correct toe burden, the top of the bench may be too close to the explosive column resulting in blow outs. Pre-splitting of the face greatly improves the situation and is generally practiced wherever cast blasting takes place, for both this and other good reasons. – Young’s modulus (GPa) Range (Gpa)
Effects on Throw
Very restricted throw
8 – 30
Acceptable movement possible
Movement easily achieved
Young’s Modulus determines the extent to which the available explosives energy is divided between: “Shock” energy (responsible more for fragmentation) and “Gas” energy (which is more responsible for (heave) movement). Low modulus rocks, which also tend to be weaker, absorb much of the available energy as Shock resulting in less energy remaining to propel the rock forward. This is why the blasting of mine dumps tends to be a frustrating experience. On the other hand, high modulus rocks are more elastic and resist crumbling under explosive pressure. As a result more energy is available to drive bulk rock movement. Figure 2 illustrates the effect of Young’s Modulus on face velocity. – Rock density Density determines the mass of the given volume of rock and hence the energy required to move it. Other things being equal (which they never are) face velocity is likely to be proportional to rock density. Thus high density rocks tend to require higher powder factors.
Figure 4 Highwall above coal seam
– Bench height This is normally controlled either by geological factors or safety considerations. Figure 3 illustrates why high benches throw better. High benches create more spoil
Explosives Today - Series 4, No 5
Figure 5 Comparison of the expected throw for angled holes compared to what is achieved with vertical holes.
constraint hence increased rehandle cost. • Height adds to the trajectory of thrown material • The “drag” effects of the uncharged collar and bottom contact have a relatively small effect over the whole face • High benches already have a significant proportion of their faces in tension, due to a lack of lateral constraint and the thrust of the blast is therefore “working with nature” The definition of adequate bench height for good movement is usually expressed in terms of the ratio of bench height to burden as follows:
The size of the dragline or shovel has an impact on the final bench width design.
Coal seam thickness and dip The thrown rock is able to travel further with higher coals seams because of the additional height of the flight before
1.5 – 2.5
– Bench width The greater the bench width, the further must the back holes move their burdens to clear the coal seam and land where they do not have to be handled by the dragline. In addition, the back rows of holes tend to suffer from progressively more choked conditions and are less able to move the
In Southern Africa the midburden comprises mostly of sandstone with an average of 2m layers of shale on top of the seam. Overburden comprises of mostly floating sandstone with softs included. Overburden coal seam is usually called Upper 4 Seam, which is soft. To protect against coal loss on the roof it is advisable to drill 1m short and back fill the holes.
ROCK DATA Medium grain Sandstone Density 2.54 g/cc E = 28 GPa µ = 0.2
landing. The same holds for seams dipping down from the face, whereas less throw can be expected from seams dipping up from the face.
Face Velicity (m/s)
rock. As a rule, bench widths in excess of 5 times the reduced burdens tend to yield reduced throw, whereas up to three times the reduced burden will yield good throw. (Reduced burden is the square root of the product of the burden and spacing.)
14 13 12 11 10
Hole diameter = 251mm Explosive = Anfex (0.8 g/cc) 6 x 15
7 x 13 8 x 11 Drilling Pattern (Burden & Spacing)
Figure 6 Effect of burden on face velocity for a constant powder factor
9 x 10
Explosives Today - Series 4, No 5
Figure 7 Problems at the front row of holes
Midburden benches mostly comprises of hard sandstone with a relative density of 2.5 and the seams are harder. Drilling 1m above coal will leave shale layers with an average depth of 1.5m , which can be ripped by dozers to protect roof coal losses.
Blast hole diameter Hole diameter is a basic input to all blast design and constitutes a major decision at the commencement of operations. Its effect on throw is experienced through the geometric influences of burden, bench height and stemming and is to be discussed under these headings. However as a general rule for cast blasting operations, hole diameter should be chosen such that: a) The ratio of bench height to burden will normally be at least 2:1. Cast Blasting is thus more efficiently achieved in low benches (less than 12m) by adopting smaller diameter holes b) Drilling inaccuracy can be limited to within 2.5% of the planned burden at the toe of the hole. This means that larger diameter holes, with their wider drilling patterns, should be adopted for higher benches
Since the drilling accuracy is the more critical side of the above parameters, it is better to err on the side of too large than too small a hole diameter.
Drilling angle There is a gain in percentage cast when blastholes are angled so as to impart a lifting motion to the burden. This is well known in quarry blasting where angled holes are fairly common. Although mechanical considerations have normally ruled out angled drilling where high mast rotary drills are used, the benefits of cast blasting are so positive in certain applications that there has been a significant swing to this practice internationally. Figure 4 shows the 32m high pre-split face of a cast blasting operation where 270mm holes are drilled at 75o to optimise mining profitability through improved cast over. Figure 5 shows a comparison of the expected throw for such a layout compared to what is achieved with vertical holes. The effect is particularly desirable when blasting “against the dip” of an inclined seam, but it should be remembered that drilling holes at an angle effectively reduces the burden and increases the powder factor over what is achieved with vertical holes on the same nominal pattern.
Another benefit of angled holes is that they enable the correct burden to be safely achieved at the toe without having to collar their holes at minimal standoff distances from the crest. In addition coal losses due to drag will be reduced thanks to the lifting action. The potential gains must be offset against the risk of the drilling accuracy of angled holes.
2. The variable parameters We have introduced the subject of cast blasting and described those parameters which would affect the amount of rock material cast, but over which little or no control can be exercised from blast to blast. The factors which favour casting are: • Competent rock with a pre-split face o Pre split blasted prior to infill holes, makes it easier to batter the highwall with the dragline /bucket shovel • High narrow benches • Competent rock with a pre-split face • High narrow benches • Thick coal seams dipping either flatly or down away from the bench • As large a hole diameter as is compatible with the bench height • Drilling angle less than ninety degrees