Monday, September 20, 2010

Explosives and Drilling

1. Introduction
Explosives and drilling are combined into one chapter because blasting in hard rock mines normally involves placing explosives into boreholes (blastholes) drilled in rock. Commercial explosives are mixtures of chemical compounds in solid or liquid form. 

Detonation transforms the compounds into other products, mostly gaseous. This action is exothermic, producing heat that further expands the gases and causes them to exert enormous pressure in a blasthole in addition to producing a stress wave. 

The combination of these two effects (borehole pressure and detonation wave) breaks the rock surrounding a blasthole. From its crude beginning centuries ago, blasting has become an advanced science that is now far more comprehensive than can be adequately addressed in this handbook. Practice in hard rock mines is continually changing due to technical advances as well as increasing concern with safety, security, and the environment.

This chapter is intended to provide some useful data for the hard rock miner while avoiding details of implementation. The information provided in the text is limited to certain aspects of blasting that are of particular concern in today’s hard rock mines. These include crater blasting, sulfur blasts, drilling equipment and drill patterns. The reader desiring additional information may refer to Gary Hemphill’s book, “Blasting Operations,” McGraw Hill, ISBN 0-07-028093-2 and valuable material published by the various powder companies as a service the mining industry. In conformity with the format established for the rest of this handbook, sections devoted to rules of thumb and tricks of the trade precede the text of this chapter.

2. Rules of Thumb
Powder Consumption
• Listed below is typical powder consumption in hard rock.
− Shaft Sinking – 2.5 Lb./short ton broken
− Drifting – 1.8 Lb./short ton broken
− Raising – 1.5 Lb./short ton broken
− Slashing – 0.8 Lb./short ton broken
− Shrink Stope – 0.5 Lb./short ton broken
− O/H Cut and Fill – 0.5 Lb./short ton broken
− Bulk Mining – 0.4 Lb./short ton broken
− Block Cave u/c – 0.1 Lb./short ton to be caved
− Open Pit Cut – 0.9 Lb./short ton broken
− Open Pit Bench – 0.6 Lb./short ton broken
Various Sources

Explosive Choice
• The strength of pure ammonium nitrate (AN) is only about one-third as great as that of an oxygen balanced mixture with fuel oil (ANFO). Source: Dr. Melvin Cook

Blasting Strength
• Blasting strength is a direct function of density, other things being equal. Typical explosives for dry ground (ANFO) may have a blasthole density (specific gravity) of 0.8 to 1.3, while for wet ground (slurry or emulsion) it varies from 1.1 to 1.3. Developments in explosive technology make it possible to choose any density desired, within the given ranges. Source: Dr. Nenad Djordjevic

Spacing and Burden
• For hard rock open pits or backfill rock quarries, the burden between rows can vary from 25 to 40 blasthole diameters. Spacing between holes in a row can vary between 25 and 80 blasthole diameters. Source: Dr. Nenad Djordjevic
• To obtain optimum fragmentation and minimum overbreak for hard rock open pits or backfill rock quarries, the burden should be about one-third the depth of holes drilled in the bench. Source: Dr. Gary Hemphill
• To obtain optimum fragmentation and minimum overbreak for stripping hard rock open pits or quarrying rock fill, the burden should be about 25 times the bench blasthole diameter for ANFO and about 30 times the blasthole diameter for high explosives. Source: Dr. Gary Hemphill
• The burden required in an open pit operation is 25 times the hole diameter for hard rock, and the ratio is 30:1 and 35:1 for medium and soft rock, respectively. The spacing is 1 to 1.5 times the burden and the timing is a minimum of 5 ms (millisecond) per foot of burden. Source: John Bolger
• The burden and spacing required in the permafrost zones of the Arctic is 10-15% less than normal. Source: Dr. Ken Watson
• When “smooth wall” blasting techniques are employed underground, the accepted standard spacing between the trim (perimeter) holes is 15-16 times the hole diameter and the charge in perimeter holes is 1/3 that of the regular blastholes. The burden between breast holes and trim holes is 1.25 times the spacing between trim holes. Source: M. Sutherland Collar Stemming
• The depth of collar for a blasthole in an open pit or quarry is 0.7 times the burden. Source: John Bolger
• The depth of collar stemming is 20-30 times the borehole diameter. Source: Dr. Nenad Djordjevic
• For open pits or back-fill rock quarries, pea gravel of a size equal to 1/17 the diameter of the blasthole should be employed for collar stemming (i.e. ½ inch pea gravel for an 8½-inch diameter hole). Source: Dr. Gary Hemphill Relief Holes
• Using a single relief hole in the burn cut, the length of round that can be pulled in a lateral heading is 3 feet for each inch diameter of the relief hole. For example, a 24-foot round can be pulled with an 8-inch diameter relief hole. Source: Karl-Fredrik Lautman
• It has been found that a relief hole of 250 mm (10 inches) will provide excellent results for drift rounds up to about 9.1m (30 feet) in length. Source: Bob Dengler Blastholes
• The cost of drilling blastholes underground is about four times the cost of loading and blasting them with ANFO. Present practice is usually based on the historical use of high explosives where the costs were about equal. An opportunity exists for savings in cost and time for lateral headings greater than 12 feet by 12 feet in cross-section by drilling the blastholes to a slightly larger diameter than is customary. Source: Jack de la Vergne
• The “subdrill” (over-drill) for blastholes in open pits is 0.3 times the burden in hard rock and 0.2 times the burden in medium/soft rock. Source: John Bolger
• “Sub-grade” (over-drill) is in the order of 8 to 12 blasthole diameters. Source: Dr. Nenad Djordjevic

Ground Vibration
• The ground vibration produced by the first delay in a burn cut round is up to five times higher than that generated by subsequent delays well away from the cut. Source: Tim Hagan

Crater Blasting
• Crater blasting will be initiated if the charge acts as a sphere, which in turn requires the length of a decked charge in the blasthole to be no more than six times its diameter. Source: Mining Congress Journal Labor Cost
• The labor cost for secondary blasting can be expressed as a percentage of the labor cost for primary mucking. For Sub-Level Cave and Crater Blasthole stoping, it is around 30%; for Sub- Level Retreat it is closer to 10%. Source: Geoff Fong

• Percussion drilling is required for drilling blastholes in rocks with a hardness of 4 or greater on the Mohs’ scale (Refer to Chapter 1). These are mainly the volcanic rocks. Rotary drilling is satisfactory for softer rocks, mainly sedimentary. Source: Dr. Gary Hemphill

3. Tricks of the Trade
• To achieve optimum fragmentation in hard rock mining, the explosive of choice is usually that with the highest detonation velocity and the maximum available energy density. Source: Dr. Melvin Cook
• The way to have your blasting pattern designed for a particular application is to call your friendly local powder agent to do it for you. This service is usually performed promptly and without charge. Source: Reid Watson
• The way to teach new hands to drive a burn cut drift is to start with a short round. Have them drill and blast only the cut until they do it correctly. Then, let them drill and blast a whole round at once. When they have mastered the short round, they are ready to try a full round. Source: Marshall Hamilton
• The average fragment size increases and the uniformity of fragmentation decreases when the depth of charge (burial distance) increases. Source: Liu and Katsabanis
• If the length of blastholes is equal to that of the relief holes in a burn cut round, the free face for the explosive at the toe of the blastholes is limited. If the relief holes are drilled a foot longer than the blastholes, the result will be less bootleg. Source: Tim Hagan
• The accuracy of delay timing in commercially available caps is sometimes insufficient to ensure proper sequence. A simple way to overcome this problem is to use only alternate delays (i.e. 0, 2, 4, 6, etc. instead of 0, 1, 2, 3, 4, etc.). Source: Tim Hagan
• Inserting a plastic sleeve right after drilling a blasthole in permafrost overburden will prevent it from becoming choked with ice build up before it can be loaded with explosives. The procedure is also beneficial in some bad ground conditions. Source: Jim Tucker
• Button bits normally give higher penetration rates but are more prone to deviation in long holes than cross bits. Source: Tamrock
• For long hole drilling, rod life, thread reversing, and redistributing the rods in the drill string considerably prolong wear life. Source: Tamrock
• Any surface concrete structure designed for a new mine (or added to an existing mine) should include plastic pipe inserts suitable for loading explosives to facilitate ultimate demolition.
Source: Peter R. Jones
• On surface, the contract specifications may require a delay of seven days before blasting near freshly poured concrete because the set may be disturbed by vibration. This is not the case underground when concrete is poured against hard rock. In many cases, nearby blasting has been carried out within 6-8 hours of a concrete pour with no ill effect. Source: Jack de la Vergne
• Steel reinforcing (rebar) can be salvaged from concrete members being demolished by drilling short holes and setting off small charges with the same delay. This procedure vibrates the steel, which cracks the cover. The cover concrete is then easily scaled off to reveal the rebar cage. Source: John Newman

4. Explosive Selection and Types
Following are the main criteria applied to select an explosive for a given type of blasting.
• Available energy per unit weight of explosive
• Density of the explosive
• Detonation velocity
• Reaction rate

5. Explosive Types
Following are the more common explosives used in the hard rock mining industry.
ANFO is the most prevalent explosive used in the mining industry because it is the least expensive and the safest to transport and handle. Standard ANFO is defined as a mixture of prilled AN having 1% inert coating and 5.7% No.2 diesel Fuel Oil as a reducing agent. This combination results in a product with a density of 0.84 and an oxygen balance of –0.5% (½% by weight oxygen deficient).

Typical operating densities range from 0.8 to 1.2 g/cm3. Packing or crushing AN prills will alter density (and sensitivity). Energy output and sensitivity are both affected by the oxygen balance (stoichiometry) of the mixture. ANFO type explosives are susceptible to water and, therefore, not suitable for wet blastholes. ANFO explosives may also pose an environmental dilemma resulting from their high nitrate content.

Slurries (Water Gels)
Due to the susceptibility of ANFO products to water, slurries were developed to replace ANFO in wet conditions. AN was dissolved in water and mixed with a fuel (in this case another oxygen deficient explosive, such as TNT) and surrounded by gum to produce a water-resistant explosive. The result was a product with higher bulk strength than ANFO suitable for use in wet conditions. The disadvantages of slurries include higher cost, unreliable performance, and deterioration with prolonged storage.

Emulsions were developed to overcome the main disadvantages of slurries. Emulsions consist of a continuous (oxidizer) phase and a discontinuous (fuel) phase combined using an emulsifier and a bulking agent, such as micro-balloons. They can be kicked-up with powdered aluminum. Emulsions have high energy, reliable performance, resistance to water, and relative insensitivity to temperature changes. The direct cost of an emulsion explosive is higher but this is offset by time saved in loading and a reduction in nitrate content of broken muck.

Dynamites are explosive mixtures made of nitroglycerin made stable by dissolving it in an inert bulking agent. The employment of dynamites has greatly diminished because of high cost and higher risk in transport and handling. Dynamites are still used for certain applications, such as shaft sinking with hand-held machines.

6. Crater Blasting
Crater blasting techniques (VCR@, MVCR, HCR, and HRM) are often employed for larger underground ore bodies in hard rock mines stoped with bulk mining methods. (Refer to Chapter 3 – Mining Methods for additional information.) Crater blasting has been used successfully for drop raises and even for shaft sinking. No real success has been achieved using crater blasting for lateral headings.

Crater blasting is based on the theory that a short stubby (spherical) powder charge will break more hard rock than a long thin cylindrical charge of the same weight. A spherical charge enables better use of the detonation wave than the cylindrical. The cylindrical charge mainly depends on the expansion of gases generated by the explosion (bubble effect). In theory, the volume of rock shattered by the detonation wave is proportional to the square of the detonation velocity. Based on this theory, explosives with high detonation velocities are employed, such as dense slurries that may have a detonation velocity 50% higher than normal.

Any blasting is more effective when it takes full advantage of gravity. Therefore, the blasting method will be more efficient if the back is blasted down in horizontal slices to the opening below. Following this procedure eliminates the requirement for a slot raise and best allows for the practice of leaving broken ore in the stope while pulling only the swell (“deferred pull”). The technique is often modified to employ a slot raise and take blocks or vertical slices with each blast instead of the classic horizontal slice. The vertical slice means blastholes are only used once; therefore, less likely to require re-drilling. Vertical slices also avoid the problem of blasting the last horizontal slice. In theory, the vertical slice provides both a horizontal and vertical free face. Experiments at the Battelle Institute indicate that a larger crater results when a second face is available. A disadvantage to vertical slicing is that the broken ore next to the fresh face must be pulled clean before the next round is blasted. This makes it difficult to leave broken ore in the stope for support (i.e. “deferred pull”).

Employing the indexing theory may provide further efficiency. The indexing theory postulates that the common crater created by two or more adjacent craters is larger than the sum of separate, single crater blasts. This indicates that the method will be most effective if a whole slice or block is blasted at once.

7. Sulfur Blasts
A sulfur blast is a subsequent air blast detonated by the combustion of sulfide dust produced from an underground blast. Sulfur blasts typically produce large quantities of noxious sulfur dioxide gas. Sulfur blasts are not uncommon in base metal mines and may be more severe where crater blast techniques are employed with large diameter blastholes. They are more frequent in development headings where the use of millisecond delays is not practical. In theory, the flame from late delays ignites the cloud of fine dust particles from initial delays. The phenomenon is believed to occur through ignition and exothermic roasting of sulfide dust particles enhanced by the addition of more dust swept up by the initial propagation.

Employing the following practices can reduce the frequency and severity of sulfur blasts.
• Use explosives with a low detonation temperature
• Use explosives with good oxygen balance
• Practice fogging and washing down
• Suppress dust with lime or limestone dust (as in a coal mine)
• Provide adequate collar stemming of blastholes
• Practice stemming with limestone and decking with limestone dust in bags
• Practice popping a bag(s) of limestone dust with half a stick at zero delay
• Blast between shifts and on weekends when the mine is evacuated

8. Drilling Blastholes
To be economical, blastholes drilled in hard rock require drilling equipment capable of both rotation and percussion. Holes could be drilled with good progress using a rotary drill and tri-cone bit, but the bit life would be too low. Diamond drills (that also have only rotation) drill with good progress; however, the bit cost and the amount of energy required to be transferred to the bit (that grinds to dust rather than chips) are too high. Consequently, miners normally employ rotary-percussion drills in hard rock formations.

Two basic types of rotary-percussion drills exist: (1) top-hammer and (2) ITH. This nomenclature is based on which end of the drill string the drill is located. In general, hard rock miners use top hammer drills for holes less than 4 inches (150 mm) in diameter, and ITH drills for larger holes. Miners still use blade bits (chisel or cross) for drilling small diameter blastholes in hard rock; button bits are normally more economical in the larger diameters.

Wet drilling has a slower penetration rate than dry drilling, but underground drilling requires water for dust suppression. This is an acute problem for mines in arctic regions where a brine solution is required to avoid freezing. One arctic mine employs dry drilling underground. Efforts to duplicate the system elsewhere have so far been thwarted, mainly because of silica content in the ore.

9. Drill Patterns
The V-cut, also called “wedge,” “cone,” and “pyramid,” is typically drilled at an inclination of 30 degrees from the drive axis. The burn cut (described below) has superceded the V cut as the standard drill pattern for development headings. 

Occasionally, a good application arises for the V-cut round. An example is when the muck is “thrown” from an entry portal on a mountain or steep hillside. Benching (a type of V cut) still has application for shaft sinking, especially in bad ground. At small mines in some developing countries, the V-cut prevails for lateral headings, despite great efforts to change old habits.

Burn Cut
The burn cut is preferred because it permits longer rounds and better fragmentation. The disadvantage is that it requires careful attention in the design and execution of the drill and blast pattern. The key to a successful burn cut round is the cut itself. The first cut holes (relief holes) are not loaded. These holes and immediately surrounding loaded holes (primary holes) must be drilled precisely parallel.

In a lateral heading, it is typical to have two or three central relief holes drilled and reamed. The optimum drill and blast pattern for the cut (and the whole round) may not be predictable in advance since it is dependent upon the actual ground conditions. The pattern also depends on the miners.

Often, different crews on the same heading have separate patterns for the same round. Good miners may employ one or two reamed cut holes, while the less able may need three. In small headings (jackleg or long-tom), after the first cut hole is drilled but not yet reamed, the miner uses a loading stick to align the adjacent holes parallel. Alternatively, the miner leaves the longest steel in the hole and retracts it to line up the drill. In a large lateral heading, a drill jumbo is employed incorporating a mechanism to automatically line up the steel to drill parallel holes.

Big-Hole Burn Cut
The big-hole burn cut occasionally employed for hard rock mining was first developed in the USA and Europe. The landmark application was used in driving the Granduc tunnel in British Columbia that achieved a world record advance rate (up to 115 feet per day). At Granduc, a separate GD 133 top-hammer drill was employed to drill a single 6-inch diameter central relief hole in one pass with a bull bit. More recently, ITH drills have been employed to drill 8 and 10-inch diameter relief holes in lateral headings to achieve longer rounds. For long rounds, design and execution of the cut is even more critical.

While long rounds are obviously desirable, maintaining a cycle is equally important. A regular cycle must be achieved to ensure consistent good footage. Another problem is that drill steel tends to wander in long holes. For example, even 1½-inch (38 mm) jumbo steel can be a problem after about 20 feet (6m) in hard rock without stabilization.

In recent years, big-hole burn cuts have been applied to shaft sinking, employing ITH drills. Originally, big-hole burn cuts were used with two or three 6-inch relief holes and then with one or two 8-inch diameter holes or a single 10-inch hole. The problem of cleaning cuttings and water remaining in the large holes was overcome by drilling the holes 2 feet deeper or rifling the cuttings and water with a half stick on the first (zero) delay.

An important dimension is the radial distance, C, between the edge of a relief hole and the primary blastholes. The maximum allowable distance can be estimated using the following formula.

In which, 
D is the diameter of the relief hole
d is the diameter of the blasthole
θ is the crater angle measured in degrees
If θ is assumed at a reasoned value of 30 degrees, the equation reduces to
C = 1.43 D - 2.43 d
Determine the furthest distance a primary blasthole can be placed from a relief hole and still achieve good propagation.
1. The primary blasthole is 1½ inches in diameter
2. The relief hole in 8 inches in diameter
3. The crater angle is 30 degrees
Solution: 11.4 - 3.6 = 7.8 inches
(In practice, the primary blastholes would be collared at a distance of about 4 inches from the relief hole, to allow for deviation.)