Lesson 3 – Drilling & Blasting

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When considering drilling in a quarry environment, it is important to analyze not the present needs of the operation, but also the growth path the site will take in succeeding years.

Drilling is one of the critical elements in the drilling and blasting process. A blasthole is merely a cylindrical vehicle designed and strategically situated to hold and contain an explosive charge so that it can be detonated in the most efficient and optimum manner possible. No blasting system will be truly effective if the hole is poorly placed; i.e. excessive burden and spacing, sized incorrectly for desired results, insufficient in depth, etc.

The drilling phase is the most expensive in the drilling and blasting portion of production, requiring a sizable investment and upkeep. The impact of improper drilling can be felt throughout the remainder of the production cycle, such as excavating and hauling, crushing, screening and so on.

The purpose here is to concentrate on some of the factors that should be considered when evaluating and selecting drilling equipment for a mining project. The primary focus is on surface operations, rather than underground.

Selection of hole diameter

The first step in the process is to determine what hole size or diameter is most befitting the application, bearing in mind that this could change over time as the operation grows and matures. This is probably the most important single factor since it will in large part determine the size, quantity and type drill or drills that will be needed.

Here are some (but not necessarily all) of the factors involved to determine the optimum drill capability:

• Required production.
• Terrain.
• Material characteristics.
• Type and size of excavating and hauling equipment.
• Proximity to vibration-sensitive areas.
• Bench or “lift” height.
• Explosives type and size.

All items should be examined and considered before making a final decision.

Required production

If intended annual production is one million tons, to choose a medium-sized rotary drill rig that is capable of drilling a 7-7/8-in. (200 mm) diameter hole and producing 3 to 5 million tons per year on a single-shift operation would be excessive.

Conversely, an installation that requires 5 million tons or more per year might consider using something other than a crawler-mounted drifter drill capable of drilling only up to a 4-in. (102 mm) hole, as a large number of drills would be required. It may be, however, that with all other things being considered, such as improved blast fragmentation, this might still be the better overall choice.

Also to be considered, is the possibility of selectivity — where hole size may have to be reduced to effectively define the ore area or address a vibration issue. Another need here might be for a more maneuverable drill rig for difficult terrain or a truck-mounted unit, able to move more rapidly from one selected area to another, thereby reducing non-productive drilling time.

Terrain

The larger the hole diameter, generally the larger the drilling platform. Big drills, even those mounted on tracks are more limited in their ability to traverse adverse landscapes than smaller machines.

If you are into a pioneering phase, then you should definitely pursue a study of the gradability and “traversability” of various drill units. As the operation matures and will gradate into a smoother series of benches, the constraints will be altered. Therefore, initially it might be wise to choose a smaller hole diameter and a smaller drill(s) until such time as the development proceeds to the point where larger holes and larger drill rigs can be utilized.

If the initial development portion is of insufficient duration to justify purchasing the lighter, smaller drills, then rentals or contract drilling could be considered for the interim, then purchasing the larger rigs when full-blown production is achieved.

Material characteristics

The previously mentioned subject of selectivity under “production” has a bearing under “material” as well. However, what is referred to in this discussion is primarily the characteristics of the rock that lend themselves to drillability and blasting fragmentation.

These are some of these characteristics that influence hole size and subsequently drill selection:

• “Hardness” (compressive strength of the rock). This subject will be approached in more detail in the following sections, but basically, percussive drilling is not as radically effected by rock hardness as is rotary drilling. This is because there are certain restraints placed on rotary bit bearing loading, particularly as the diameter decreases. The other effect of rock hardness is its resistance to blast fragmentation, particularly if it is somewhat homogeneous in nature. This can effect the distribution of the explosives in the bank, so that smaller holes, closer together may have to be considered, depending on desired fragmentation.
• Rock structure or existence of joints, fractures, bedding planes or faults. If the rock, even though hard, is rather friable in structure, (such as a well-jointed basalt or “trap rock,”) then larger holes farther apart may be used and still retain optimum fragmentation. On the other hand, if a radical series of fracturing is present which causes rapid attenuation of explosives detonation waves, then smaller holes closer together may be necessary.

Another structure factor might be the presence of a blocky “cap” rock at the upper part of the hole. This can particularly occur in the “top lift” or bench nearest the weathered zone of the material. This bears consideration because the larger the blasthole diameter, the more “stemming” is required in order to better confine the explosive charge. In other words, the top of the explosive charge may be below the cap rock portion, thereby contributing nothing to its fragmentation unless “decking charges” or “satellite holes” are drilled and loaded, or secondary blasting or breaking with a hydraulic hammer or drop ball is done on the pit floor. This situation should be evaluated from the standpoint of cost savings via the larger hole versus the cost of secondary breakage of the cap rock.

Yet another structural feature of the rock that can effect hole-size determination is the tendency for it to “backbreak” excessively causing large fragments or blocks of unblasted material to be dislodged into the muckpile where secondary sizing will have to be done.

This problem can most severely occur when blasting “up-dip” parallels with the strike of the rock. This backbreak can also cause problems on drilling succeeding blasts, since it may call for angling or “looking” holes in the first row or rows in order to maintain proper toe burdens.

One of the solutions to alleviate this situation may be to either use “deck charges” in the normal stemming area of larger holes, drill smaller-diameter “satellite” holes in between the larger holes, or to drill smaller-diameter holes exclusively, where the explosives column can be brought higher, and loaded lighter, better-distributed charges can be used, thereby reducing damage and minimizing the backbreak to a manageable level.

If toe problems still persist and angle drilling is required, it should be remembered that large drills have limited capabilities in this regard (up to 30 degrees as an optional feature). Smaller drills drilling smaller diameter holes have better capabilities in angle drilling, but trade-offs are present in difficult setup and hole accuracy. Also bench heights may have to decreased. This will be discussed in more detail later.

Type/size of excavating/hauling equipment

Some are of the opinion that because large excavators and haul units are employed the blastholes should be as large as possible and the drill pattern spread, since larger sizes of blasted rock can be handled. To a certain extent this is true, but it can be argued that the primary purpose in larger excavating and hauling units is to promote greater production capabilities more economically, not to save money on the drilling and blasting phases.

Caution should be exercised not to determine blasthole size purely on the basis of this factor. On the other hand, if excavating and hauling equipment is relatively small, then careful consideration of hole size relative to desired fragmentation should be made.

Proximity to vibration-sensitive areas

Most operators who employ drilling-and-blasting techniques are well aware of the vibration restrictions imposed on them in today’s lawsuit-conscious world. Obviously the larger the hole, the more weight of explosives per hole, and unless you can employ more than one delay per hole or advanced vibration control techniques, the higher explosive charge can yield higher vibrations.

Many factors should be taken into consideration before selecting a drill. Of course, you won’t want to drill and load 12-1/4 in. (311 mm) holes adjacent to a new subdivision, but with proper delay utilization and monitoring you could address the vibration concerns with a 6-in. (152 mm) hole at a considerable cost savings over even-smaller diameter holes.

To insure proper drill selection it may be necessary to do some vibration-control research and testing, particularly in a new operation, or construction project without previous background history.

Bench lift or height

If there is an existing bench height, then selection of blasthole size has to be made with this in mind. In some areas, legislation has dictated maximum heights so that some of the selection process has been eliminated. Wall control and stability may also be a significant factor in deciding bench heights.

On the other hand if bench height is variable there is more latitude. (In many cases bench or bank height is pre-determined by production requirements and hauling and digging equipment.)

There are other considerations that need to be evaluated in this process:

• Blow energy to the drill bit from an “out-of-the-hole” drifter drill dissipates considerably as depth increases. This subject will be discussed in subsequent text, but holes drilled by this type of drill engine are in the range of 5 in. (127 mm) and below. While there is some variance in design and power, realistically 30 to 40 ft. (9 to 12 meters) is about the maximum economic depth for this range of hole.
• Percent of available hole volume for loading with explosives increases with bench height. Simply stated, this is the relationship between the volume of the explosives column to the “pay volume” or drilled volume of the hole above the floor, (without subdrilling), stated as a percent.

At the outset, it was stated that a blasthole is a cylinder whose prime purpose is to efficiently accommodate an explosives charge. A blasthole cannot be loaded right up to the top of the ground with explosives, since there is no confinement of the energy to break the rock and the hole will “rifle” or “gun-barrel” throwing fly rock creating a safety concern, high air-over pressures and poor fragmentation.

Therefore, it is necessary to maintain a specific depth from the top of the hole to the top of the explosives column, and fill this portion of the hole with an inert material such as small angular crushed stone (stemming). Larger diameter bore holes will require larger crushed stone to insure the blast will be confined for optimum performance and safety.

Since the detonation of an explosives column tends to direct its energy toward the path of least resistance, which most often is up and forwards, it is generally necessary to “sub drill” or drill below the floor or final grade in order to place explosives to insure the blast breaks to grade. When detonated the explosives in the sub drill will help to create a shear plane and displace the blasted material at or below the desired floor level.

(In some instances such as surface coal or horizontally bedded sedimentary rocks, this may not be necessary or desirable.)

Since larger diameter holes are generally placed farther apart, the subdrilling is somewhat greater in order to break between the holes and not cause “high-bottom.” Therefore the length of the explosives column is the bench height plus the subdrilling less the stemming. However, the “pay” portion of the hole is only the length corresponding to the bench height. A relationship can then be set up showing the ratio or percent of the volume of the explosives column to the pay volume of the hole. If the explosive used is fully “coupled,” or fills the complete cross section of the hole, as with poured or pumped explosives, then this relationship becomes merely the ratio of the length of explosives column to pay length of the hole.

Explosives type and size

Finally, another factor that should be taken into consideration is explosives type, size and method of loading. Explosives generally take on more favorable properties of detonation in larger holes and are usually easier to load, especially with the advent of bulk loading systems. Some cost advantages may prevail that are worthy of attention.

Drilling methods

The variables discussed cover what should be considered in selecting a blasthole diameter and also along with it, an approximate bench height. It has also stressed how critical this is in determining the quantity, type and size of drilling equipment that is required. Actually, there is an intermediate step required here before final determination; and that is an evaluation of what general drilling method (or methods), will be best suited for your particular application.

Drilling methods, herein referring to the actual mechanics of the process, can be broken down into two broad categories, and then further divided into subsections as follows:

Percussion:

• Drifter or out-of-the-hole drills (OHD).
• Air-actuated.
• Hydraulic fluid-actuated.
• Down-the-hole-drills (DHD)

Rotary:

• Blade (or drag) bit rotary drilling.
• Roller-cone bit rotary drilling.

Percussion drilling

This method of drilling is, in simple terms, nothing more than mechanization of the venerable “hand steel and hammer” method used in mining 100 years ago and more, wherein the end of an integrated piece of steel and bit were struck by a hand-held hammer, and then rotated by hand between each blow in order to re-index the bit so that it didn’t “rifle.”

The “ingredients” in present-day percussive drilling technology are still the same:

• A bit in order to actually penetrate the rock and cause it to fail.
• A method of imparting a sharp blow that can be transmitted to the bit.
• A rotational device so that the bit can be incrementally indexed in the hole.

In addition to the above, a means of removing or flushing the drill cuttings and dust from the hole must be provided. The drill bit must, by necessity, be highly resistant and hardened in order to absorb the blow energy and penetrate the rock. Today’s bits are generally detachable and replaceable and have cutting edges of a very hard alloy, such as tungsten carbide in the form of inserts or buttons.

The blow is delivered by a piston reciprocating within an enclosed cylinder. Rotation can be achieved by either a mechanism contained within the drill engine or an external device, such as the rotation or power head on a rotary drill. Hole flushing is generally performed using compressed air.

Drifter drills (OHD or top hammer)

In this method of drilling, the drifter, or drill engine, moves along a guide or track on the drill carrier. It contains both the piston and the rotation components. Piston actuation can be achieved by either compressed air or hydraulic action, (the merits of each will be briefly discussed later). The piston imparts its energy to a shank piece or striking bar, inserted in a chuck, located at the bottom of the drifter. This shank passes the blow along to the bit via a series of drill “steels” or rods connected by detachable couplings.

Within the drifter, there is a “blow tube” that directs compressed flushing air down through a drilled hole in the piston and drill steel to drilled passages in the bit that force the drill cuttings out the annular space between the hole diameter and the drill steel to the surface. In addition to compressed air alone, water, foam or some type of “detergent” can be combined with the flushing air to cause cuttings and dust to adhere, thereby controlling airborne emissions.

The drifter and drill “string” are fed up and down the drill guide by (generally) a motor and chain device, in order to keep down-pressure on the bit and also to allow the addition and removal of drill steel. Air-operated drifter drills, generally use air at pressures of about 100 to 125 psig (690 kPa to 862 kPa). To use higher pressures causes excessive wear and breakage of the drilling accessories. Hydraulic-operated drifters employ similar principles of operation, except that hydraulic fluid is relatively non-compressible.

The piston stroke is therefore much shorter than air drifters, but the work performed is compensated by greatly increased blows per minute (BPM). Hydraulic drifters still need a certain amount of compressed air in order to flush the cuttings from the hole. Drifter drills, whether air or hydraulic, are limited to the hole sizes below 5 in. (127 mm). To go higher in hole size and maintain economical production would require a much larger drill engine or drifter since pressures are limited. This in turn, means larger bits and steel systems, which poses handling problems.

Down-the-hole drills

In drifter drilling, one of the big drawbacks is that the blow energy to the bit from the “out-of-the-hole” piston, is attenuated greatly through the absorption by the drill string. The deeper the hole, the less the efficiency. A solution to this would be to put the piston actually in the hole where its energy could be applied directly to the bit. This is the principle of the downhole drill, wherein the cylinder and piston are encased in a sleeve that actually goes into the drill hole with the bit. The rotation mechanism is still retained externally. At the present, all DHDs are compressed air operated.

Note that while provision still has to be made to provide a passage of a certain portion of the air for hole cleaning, the DHD uses exhaust air after it has been applied to the piston, rather than a diverted portion of the main air supply as in drifter drilling. Also a water check valve is supplied to prevent ground water from entering should the air be shut off. The only moving part is the piston, which acts as its own valve by opening and closing the air intake and exhaust ports. It should be obvious, that in order to accommodate all these components in a confined space, the piston becomes relatively small and indeed must be considerably less in diameter than the bit.

The reduced piston size reflects two very important aspects of DHD drilling:

• In order to maintain ample blow energy to the bit with a piston of less area of applied pressure and smaller mass, higher pressure compressed air must be used. Generally, this is 250 psig (1723 kPa) nominal pressure at the compressor receiver tank and above, (to about 350 psig or 2412 kPa at present).
• The minimum hole size is necessarily limited due to the space requirements and the metallurgy of the components, particularly the piston, to withstand the forces involved. For blasthole purposes, and to preserve the economics of production, this minimum hole diameter would be about 4-1/2 to 5 in. (114 to 127 mm).

Downhole drills have one other consideration in their favor, and that is since the blow is produced in the hole, airborne noise is reduced considerably over drifter drills. Percussion drilling, whether by OHD drifter or DHD downhole drill, is possible in all types of material regardless of hardness, strength or abrasion.

Rotary drilling

The concept of rotary drilling as a method differs entirely from percussive drilling. In the latter, the rock is caused to fail by imparting a sharp blow onto a bit and using a rotation device to merely turn the bit a sufficient amount to keep it indexed to cut a circular path. This rotation in percussive drilling is best kept to a minimal value, and feed or down-pressure should just be sufficient to keep the bit fed to the bottom of the hole.

In rotary drilling, no blow is struck, and the rock is made to fail by a combination of down pressure and rotation speed. In blasthole work, the cuttings are flushed from the hole also by compressed air, but consideration here is more toward providing enough volume to maintain a suitable “bailing velocity” rather than pressure. Actual pressures in rotary drilling run in the range of 50 to 100 psig (345 to 689 kPa).

Remember this, as it is a very important factor that will come into play later on when we get to selecting equipment. Rotary drilling then, is more of a brute-strength operation and requires massive pulldown systems and raw rotation power. However, with a suitable high-pressure air supply, a rotary drilling machine can be used with downhole drills should conditions dictate. Since the principles are different, it would stand to reason that bit design for rotary drilling would have to differ considerably over that of percussion.

For blasthole rotary drilling, bits take usually two configurations:

• Drag or blade bits.
• Roller-cone bits.

Drag or blade bits

The “drag” or blade bit has fixed “wings” that penetrate the rock and “gouge” it out. Some of these bits are supplied with replaceable blades. For harder materials, the blades are tipped with tungsten carbide for longer wear life.

These bits are generally confined to quite soft formations (usually less than 20,000 psi compressive strength) and in smaller diameters in the range of 3-1/2 in. to 6 in. (89 to 152 mm). Their rigid design prohibits the metallurgy from withstanding the higher pulldown and rotation stresses necessary in harder formations and larger diameters.

Roller-cone rotary bits

Roller-cone (or sometimes referred to as “Tri-Cone” or “Rolling Cone”) bits are the most utilized for blasthole drilling. The bit actually consists of three conical shaped members that rotate on a combination of roller and ball bearings. On these cones are mounted some kind of cutting tooth design depending on the hardness and compressive strength of the material being drilled.

On some of these bits the teeth are cut on the cone body itself for softer formations, whereas for harder rock, the cutters are hardened steel or tungsten carbide inserts. The tooth patterns on adjacent cones complement each other so that they will contact different portions of the rock. On some bit designs, the axes of rotation of each cone is offset from the others to impart a “scraping” action to the material to be drilled.

The most critical point of any roller-cone rotary bit is in the bearings. In addition to failure due to excessive pulldown or bit loading (this will be discussed later), these bearings need to be kept cool and clean in an otherwise hostile environment. Previously, we mentioned the need for sufficient compressed air to be delivered through the drill pipe to the bit to properly flush the cuttings from the hole with a suitable bailing velocity. In addition, a portion of this air needs to be diverted through passages in the bit body to clean and cool the bearings.

While we did say that air pressure was of less importance in rotary drilling, it is vital that sufficient pressure be delivered to the bit in order to adequately perform the cooling-cleaning function.

Curiously, when rating rotary drilling rigs, the emphasis is placed on the “pulldown” power in terms of pounds or kilograms of force exerted by the feed system. In reality, most of the power consumption in a rotary drill rig is apportioned off in driving the air compressor (about 60 percent) and in rotation power (or torque to maintain rpm under heavy resistance) (about 25 percent), whereas “pulldown” consumes less than 5 percent.

Rotary-bit design weight

Another vital consideration in selecting a rotary blasthole rig, is the subject of rotary bit design weight. This is a function of the amount of bit loading that the roller-cone bit bearings can take before failure, as mentioned in previous text. This is assuming that adequate cooling and cleaning air is present. The larger the bit diameter, the larger the bearings, and all else being equal, the more weight they can sustain.

To illustrate this, let us assume that we have selected a hole diameter of between 184 to 200 mm (7-3/8 to 7-7/8 in.), and have decided that rotary drilling with roller-cone bits is the way to proceed. It should be obvious that we wouldn’t want to choose a rotary drill capable of exerting bit weights of 60,000 lb. (27,216 kg) as it would be too much machine for the application, and we would have to hold back on the pulldown in order to keep from tearing up the bearings in the bit.

On the other hand, selecting a rotary rig capable of only 30,000 lb. (13,600 kg) would not allow us maximum efficiency and productivity out of the hole size. This factor of bit loading also bears another important consideration in our drill selection. Since as bit size decreases, the amount of allowable loading also decreases, smaller pulldown forces are necessary.

Penetration rate has pulldown as one of its variables. Therefore, rotary drilling is not as effective as the diameter decreases and the rock hardness increases. In this region, it may be possible that DHD drilling may be more productive. Remember, however, that DHDs require higher inlet air pressures, so that rig selection will have to take this into account.

Drill rigs and mountings

Up to this point, we have looked into the optimum hole diameter for our operation and have also related the various drilling methods toward that hole size.

The next step is to investigate and select the type of drill and drill mounting that would be best suited for maximum overall production at minimum cost. By going through the previous steps, in effect we have greatly narrowed down the choices.

In addition to just penetration rate, we need to view the drill rig and mounting also from the following aspects:

1. Mobility: Can it effectively move to and traverse the drilling area or areas with a minimum amount of difficulty and time consumed?

2. Simplicity: Is it a machine that can be operated effectively without undergoing extensive training and is its functioning such that operator fatigue will not be a problem?

3. Reliability and mechanical availability: Is it a rig that will not pose severe maintenance and repair problems? Are repair parts readily available and is technical assistance obtainable if necessary?

4. Productivity: Can it meet the necessary production schedule? This is “overall productivity” considering not just the mechanical availability but the time consumed in the “non-drilling” functions, such as moving from hole to hole, changing drill steel or pipe, pulling out of the hole etc.

5. Safety: Is the rig safe to operate?

Based on selection of hole diameter and drilling method, following are some areas for consideration:

Percussion drill mountings

“Drifter” or “out-of-the-hole” drills: Both the air-actuated and hydraulically actuated drifter drills generally are only available in a crawler mounting so that generally terrain mobility is not a problem. However, since grousers and pads, sprockets and rollers etc. are susceptible to wear and tear when tramming over long distances, and are of course slow, it is recommended that these rigs be transported on some form of truck or trailer in such instances. Hydraulic drills only need a small air compressor for “blow” air, and this is usually self-contained. Air drills, on the other hand, need a much larger compressor (600 cfm or 17 M3/min or larger), which is a separate unit and requires hose and/or piping to connect to the drill. Hydraulic drills, while usually heavier, are generally easier to move as a package than the air crawlers. However, if the air crawler compressor supply, can be centralized, then the drill itself is more mobile than the hydraulic version.

Drifter-type drill rigs, while fairly easy to operate, have been historically quite labor intensive, requiring manual steel handling and a lot of “bull work” in general. Also, the operator has always been exposed to the noise of the drifter and the drill dust created. Lately, though, especially in the hydraulic models, more sophisticated rod handling methods have been developed, and in many cases the operator can function from a sound-attenuated, climatized cab rather than be exposed to dust, noise and the elements.

Since the greatest enemy of hydraulic systems is dirt and dust, and drilling is the worst environment to provide these elements, the hydraulic drifter drills have up to now been the more susceptible to “down-time.” Also, repair of a hydraulic drifter must be done in a near antiseptic atmosphere, whereas, air drifters can be disassembled on the site, and reassembled with little more than wiping off the grime and re-oiling the parts. Better sealing and improved technology, have increased the “up-time” of the hydraulic drills, but still a rigid schedule of regular inspection and overhauls is highly recommended. Vibration is a problem with drifter type drill rigs. They therefore have a reduced useful life and require more maintenance than DHD or rotary drills.

If the “downside” of hydraulic drifters is the maintenance, the “upside” is that generally they drill faster than their air cousins. Evaluating one factor versus the other is going to be a major element on deciding on air or hydraulic. Both types of drifters come usually equipped to drill angle holes. One other factor should also be considered if the material is generally seamy and broken. Air drifters can effectively “hammer” themselves out of a hole if stuck, hydraulic drifters are less effective and usually need an additional “reverse-percussion” attachment.

Drilling machines are inherently dangerous just from their nature. Legs can be run over, fingers and hands lost in changing rods, clothing and flesh caught in moving chains etc. Training and awareness are the best sources for the prevention of accidents along with choosing a rig that is the product of a reliable manufacturer wherein experience has dictated the incorporation of as much safety as possible.

Compressed air

Some DHD drilling can be performed using modified smaller crawler mountings such as would normally be associated with drifter drills. A rotary head is added as is a drill pipe handling system. Usually drill pipe is only 7 to 10 ft. (2.1 to 3.05 m) in length to facilitate this. Of course, to be effective, a source of high-pressure compressed air is also necessary. Since pipe handling is more frequent and consumes more time, the overall productivity is somewhat less on this style of mounting as opposed to larger rotary drill models which are better set up with more automated handling systems and longer drill pipe lengths.

Most DHD drilling in blasthole applications is performed using larger rotary drill rigs, modified to provide the necessary high-pressure air. Therefore, the discussion of characteristics of drill rigs and mountings for this method of drilling is best covered under the “rotary drill” section.

Rotary drill mountings

Blade (or “drag”) bit rotary drilling: This type of rotary drilling is limited to about 3-1/2 in. – 6 in. (89 to 152 mm) in soft to medium formations only. The ideal drill rig for this application is a lightweight rotary drill with less than 25,000 lb. (11,300 kg) of pulldown, preferably with single-pass capabilities (no addition of drill pipe). By using a table drive rather than a tophead drive rotation mechanism, a lighter weight tower can be used, thereby lowering the center of gravity and making moving from setup to setup faster.

Roller cone bit (and DHD) drill mounting: Again in review, you will recall that the size hole involved is fairly large, and (other than the instances mentioned where DHDs are mounted on “Drifter-type” crawler mountings) a drill mounting for this type of application has to be fairly substantial.

The DHD drilling doesn’t require the pulldown and rotation forces of roller cone drilling, but generally drill pipe is longer and heavier and a substantial tower or derrick is necessary to facilitate handling. Remember, that with the inclusion of high-pressure air, both DHD and rotary drilling can be accomplished with the same drill in most instances.

Drill mountings fall into two categories:

• Truck-mounted.
• Track-mounted.

Truck-mounted

This type of mounting is really an adaptation of a waterwell rig. Its obvious biggest advantage is speed of mobility between sites and drill areas. Its disadvantages are:

• It can’t traverse severely adverse terrain.
• It does not provide as substantial and as heavy a platform for straight rotary drilling as track machines.
• It usually has two engines to maintain.
• Tire maintenance can be a problem.
• It takes longer to set up on a hole and usually requires an additional person to spot holes; (note: remote propel systems are available, but there is a definite hazard encountered here when control over a truck is performed from the rear of the machine.)

Track-mounted

This configuration overcomes most of the disadvantages of the truck mounting except it is less adaptable for rapid deployment between drilling areas or job sites.

Both rotary drill-mounting types are limited in their ability to drill angle blastholes. This capability is purchased as an option and is usually limited to 20- to 30-degree holes in 5-degree increments and in one plane only. This causes increased set up time because the rig has to approach the hole location in a plane perpendicular to the face rather than along a parallel plane as with vertical holes. This factor has to be offset against the advantages of angle drilling particularly in cast blasting (or blast casting) applications in surface coal mines.

Choice of rotary drill mountings

First and foremost it is assumed that the drill will be of sufficient size and power to handle the designated hole diameter. The air plant has to be of sufficient pressure and capacity to meet all requirements, including bailing velocities and DHD needs.

Secondly, when choosing between a truck or track mounted rotary drill rig, the evaluation has to be made as to the importance of the mobility factor versus the possibility of less maintenance and probably more efficient drilling in the sense that while on the shot, the track rig should consume less time in moving from hole to hole, thereby reducing the “non-drilling” time of the cycle.

Overall productivity as a function of cost

In the course of this chapter, we have referred to “overall productivity”. What this really means is the time spent actually drilling (“bottom-hole time”) as related to the time spent in mechanical down-time and time spent in the non-drilling functions which include moving, setting up, adding drill pipe or steel, pulling rod or steel from the hole, time “stuck” in the hole etc.

Choosing a drill rig only on the basis of its ability to “punch” down holes is therefore not enough without consideration of these other parameters.

Conclusion

Evaluating and selecting a blasthole drill can be a very involved process. Often it is not provided with the consideration that it deserves. This can sometimes be unfortunate, as the success of an entire project or operation can hinge on the ability of a drill to produce the quantity and proper sizing of the raw material. No amount of explosives expertise and technology can overcome a hastily and ill-informed selection of hole diameter, drilling method and drill rig.

BLASTING

Blasting is an area that has come under very strict scrutiny in the post-9/11 work environment. Blasting is heavily regulated and watched by federal and local agencies.

In terms of processing, blasting is the critical first step in the rock-fragmentation process. Maximum profitability depends largely on good blasting. Consider that drilling and blasting are the first operations performed in any hard-rock quarry operation. Therefore, the results of these operations will affect more down-line activities, such as loading, hauling and crushing, than any other processing operation.

Blasting should always be viewed in the “global” sense. One should examine not only the effect of changes on the drilling and blasting program, but also how the change will affect the productivity and economics of other down-line cost centers. Blasting should also be viewed in the “local” sense. No other quarry operation has more capacity to cause community dissent than blasting.

All quarry operations should have in place a proper public-relations program designed to communicate to the community that proper safety precautions and procedures are in place with regard to its blasting program.

Fragmentation

In the 1990s increasing emphasis was placed on the role of fragmentation at the operation. In particular the effect of fragmentation on crushing, load and haul, and run-of-mine leach pad efficiency has received considerable attention. Better predictive techniques have been developed, and computer-aided methods for determining the fragmentation distributions in actual blasts are now available. Fragmentation studies can lead to improved profits at many operations. For example, studies at one operation showed that the same production could be obtained with one less excavator in good digging, when compared to poor digging conditions. This is a result with both capital- and operating-cost implications.

For maximum success it is essential that the mine or quarry carefully design its blasts to achieve the desired results. These designs must be accurately implemented in the field. The blasts must be shot in a safe manner, with the area properly barricaded and all persons removed a safe distance away. Environmental affects such as ground vibration, airblast and fume production must also be controlled.

Explosives

The past 15 years or so have seen new explosive formulations reach the marketplace, and reductions in the use of some products that have been in use for longer periods. The principal newcomers have been the emulsions, and emulsion-ANFO blends usually called Heavy ANFO, that denotes its greater density than ANFO dry mixes.

Emulsions

The formulation of an emulsion is very similar to that of blasting slurries (water gels). However, the cross-linking agent used to stiffen the slurry is replaced by an emulsifying agent. This water-in-oil emulsifying agent suspends minute droplets of the ammonium nitrate (or a combination of AN with either calcium nitrate or sodium nitrate) oxidizer within the fuel. This yields a very intimate oxidizer and fuel mix that leads to high detonation velocities.

Emulsions may be bulk loaded, or used in packaged form. Packaged products are usually employed in small hole diameters. They are mechanically sensitized using microballoons to introduce sufficient air into the mix and control the density. Bulk emulsions are used in larger diameters and may be mechanically or chemically sensitized, with chemical sensitization being less costly. Bulk-loaded product fully fills the cross sectional area of the hole and delivers maximum energy to the surrounding rock. This is a characteristic of all bulk-loaded products unless they are intentionally decoupled as is often the case in presplitting. Packaged emulsion will usually result in some decoupling with a reduction in borehole pressures. This generally is not a great problem in small diameter blastholes.

Ammonium Nitrate Fuel Oil (ANFO)

ANFO remains one of the most commonly used products in quarry blasting. It is a combination of ammonium nitrate (oxidizer) and number-two fuel oil (fuel). Number-one fuel oil may be used in cold-weather applications.

Blasting grade AN prills are made by spraying molten AN into a prilling tower. Droplets fall under carefully controlled cooling conditions. The AN solidifies while falling, taking on an approximately spherical shape of relatively uniform size. Prilling tower conditions must be such that will produce a “porous” prill that will absorb the proper amount of fuel oil (6 percent by weight). For those with overseas operations especially it will be important to confirm that a porous prill is being produced. High-density prills will not properly absorb the fuel oil and blasting performance will suffer, unless these have been crushed to about -20 mesh.

Blasting-grade AN prills are typically +6, -14 mesh in size. This uniformity in the size of the prills results in a poor packing density, with considerable interstitial voids present. Hence, it is a product that typically bulk loads in a density range of 0.80 to 0.85 gm/cc. Some packaged ANFO products use a blend of sizes, where a portion of the prill is crushed, leading to densities of about 1.05 gm/cc. This product can be loaded in wet holes provided it is contained in a suitably waterproof bag.

ANFO has virtually no water resistance. Many people are of the impression that it takes several hours before water attack seriously affects ANFO. The reality, however, is that degradation of the product is immediate. Even if holes will be detonated 2 or 3 hours after loading, performance will have been seriously affected.

Therefore, whenever ANFO is to be loaded into wet holes, the blastholes should first be pumped and a plastic liner placed in the hole. The ANFO is loaded inside the liner. Care should be taken to obtain a liner that has a high integrity. Even a few pinholes are enough to allow water to attack ANFO. For hole diameters less than 5 in., using plastic liners is generally difficult. Therefore, small diameter waterproof products such as emulsions or slurries are generally used for small diameter, wet holes.

One way to increase the energy output in ANFO is to add aluminum. The reaction of ammonium nitrate with aluminum releases more energy per unit of weight. The aluminum must be of a suitable size to be reactive, but not so fine as to constitute an explosion hazard. This generally means a size range of -20, +150 mesh.

The upper limit on aluminum addition is usually about 15 percent. As more Al is added to the mix increasing percentages of the energy are trapped in a solid product of detonation. Beyond 15 percent Al by weight there is little additional energy output for the aluminum added.

Heavy ANFO

Another way to increase the energy output of ANFO is to add emulsion to it. The emulsion fills the voids between the prills, the density increases and there is more energy output per unit of blasthole volume. This class of explosives are known as Heavy ANFO. They provide a cost effective way to increase the energy output of ANFO.

Heavy ANFO may be produced solely for the purpose of increasing the energy output. However, at higher emulsion percentages by weight these products become waterproof. Such formulations can be bulk loaded into wet holes.

Experiment has shown that the performance of Heavy ANFO becomes sluggish as more emulsion is added unless the emulsion has been sensitized by gassing or microballoons. It appears that in hard-rock performance will suffer when there is more than 30 percent of unsensitized emulsion in the mix. In softer formations greater percentages of unsensitized product can usually be employed because suitable fragmentation of the rock depends to a greater degree on heave energy. The degree of non-ideal detonation introduced by the lack of sensitization means that a greater degree of the total energy is released as heave energy.

A waterproof product is typically produced at 50 percent emulsion addition. However, to obtain a product that can be pumped reliably it is common to use a waterproof Heavy ANFO containing 60 to 70 percent emulsion. Such products should always be made with a sensitized emulsion, or performance will suffer.

When waterproof heavy ANFO blend is loaded into wet holes it should always be loaded from the bottom up. This is achieved using a bulk truck with a hose that can extend to the bottom of the blasthole. The product is pumped through the hose. The hose is retracted as loading proceeds, but is always kept in the explosive. The water rises on top of the advancing column of more dense explosive. Mixing does not occur if the loading is carefully performed.

When Heavy ANFO is augured into wet holes it spatters on impact with the water, and prill goes into the solution. Water is mixed into the explosive column. Bridging may occur with portions of the explosive column separated by a water gap. Since the gap sensitivity of these products is not large this may lead to the failure of a portion of the explosive column to detonate unless it happens to be primed on both sides of the water gap.

Heavy ANFO is also produced as a packaged product. In this case it is sensitized using microballons, which improves the shelf life. Package products can be used where there are insufficient wet holes to warrant bulk loading, or in small tonnage operations. It is also used as a toe load in holes that have only a few feet of water in the bottom of the hole, and can be used in small-diameter packaged formulations.

Dynamite

There is still a considerable amount of dynamite sold annually in the U.S. However, pits and quarries have almost completely moved from the use of dynamites to small diameter, cap-sensitive emulsions and slurries for appropriate applications. Dynamites are explosive substances that depend upon nitroglycerin or nitrostarch for sensitiveness. These products are usually cap sensitive with a detonation velocity dependent upon the diameter and density.

Dynamites are used as decoupled charges in presplitting. They are also used sometimes to prime ANFO in small diameters. For this latter application a product with high detonation velocity should be chosen because it will have the higher detonation pressures (a function of the square of the detonation velocity) that are important for efficient priming of ANFO.

Explosives and blast-initiation accessories

Some of the explosive described above are cap sensitive. This means the product can be efficiently detonated by a blasting cap or delay detonator of adequate strength, or by compatible detonating cord. Small-diameter emulsions and slurries are typically cap sensitive. The manufacturer should be consulted as to the proper accessories to use. The sensitivity of some products may vary with temperature. Greater priming strength can be required when the product is to be detonated at low temperature.

Bulk-loaded explosives used in hole diameters greater than 5 in. almost always require heavier priming than a detonator alone can provide. Initiation of the bulk explosive is temperature- and pressure-dependent. Those primers yielding high detonation pressure initiate the explosive more efficiently. Thus formulations with high velocity of detonation (VOD) generally give the best results. For this reason the cast pentolite primers were developed. These generate 2.2 to 2.8 million psi detonation pressure, depending on the formulation. Various designs to provide an effective primer economically have been developed.

Cast primers are often 1.0 lb. in weight. However, primers of greater weight are also produced. These may be useful in difficult applications or with an explosive having a higher minimum primer weight. The weight per primer used in the blastholes should be 4 to 6 times the minimum primer weight. Cast primers typically have a length to diameter ration of 3:1 to 4:1. The primer should have a sufficient diameter to act on an adequate cross sectional area of the explosive charge thereby insuring efficient initiation. It must be long enough to allow the VOD in the primer to build up, providing maximum pressure off the end of the primer. Therefore, there is a trade off between length and diameter to provide effective initiation with a primer of reasonable dimensions and cost.

Cast primers are made with a single tunnel through which detonating cord can pass or with a tunnel through the primer and a cap well. The cap well accepts a down-the-hole delay detonator for in-hole delay applications

Slider primers are used for multiple priming on a single detonating cord downline. This is often used when deck loading is employed. These primers are made with a tunnel affixed to the outside of the cast primer. Detonating cord passes through the tunnel. The pigtail on the end of the delay detonator is also passed through the tunnel. Upon initiation the delay is initiated from the contact between the detonating cord and the delay pigtail. Only certain types of downlines (usually of low grain count) can be used and this information should be obtained from the manufacturer.

In some cases a stick of dynamite is used to initiate the hole. The approach is most common when priming ANFO in small diameters. The majority of dynamites do not generate the kind of pressures a cast primer provides. However, a few gelatin products detonate at very high VOD and do give high detonation pressure. When priming ANFO with dynamite this type should be used.

Detonators and initiation systems

Detonators are used to initiate the blast. These may be electronic, electric or non-electric. For modern-day blasting, delay detonators are virtually always used. Delay detonators are available for use in the hole, and also for connecting into the surface tie-in.

A delay detonator is similar to an instantaneous cap except that a delay element is included between the initiation charge that is activated by the in-coming energy, and the base charge. The delay compound burns at an accurately know rate and provides the desired delay time.

Down-the-hole delays are used alone to provide the proper firing rotation or in combination with surface delays. In the former case different delay times are used in the appropriate blastholes to provide the desired sequence of detonating holes. When used together with surface delays a constant down-the-hole delay time is often used. The in-hole delay is of sufficient duration to allow several rows of surface connections and downlines to be activated in advance of blasthole detonations. This approach avoids cutoffs and misfires that reduce blast performance and introduce subsequent safety concerns. When down-the-hole delays are used it is often possible to use longer surface delays without fear of cutoffs.

In orebodies where hot holes are possible (such as reactive sulfides) down-the-hole detonators must be used very carefully, because these are the most sensitive element to heat in the blasthole. Holes over a certain temperature are often not loaded. Top priming just before shooting is often indicated. Avoiding the use of these detonators is also an approach taken by mines where this is a severe problem.

Down-the-hole delays are often made with shock tube lead lines. These may be long lead where the shock tube extends all the way to the collar, or short lead where the shock tube is an 18- to 24-in. pigtail. These latter units are used with detonating cord downlines. The detonating cord must be compatible with the delay system used.

Surface delays provide good flexibility in blast tie-in to provide for the desired sequence of detonating holes. Delay units are made that can be spliced into detonating cord trunklines used to connect the blastholes together. Systems are also available with long shock-tube leads, eliminating the need for the more noisy detonating cord. This is especially useful for quarries because these pits are often sited in close proximity to residential and commercial areas. However, the latter systems cannot be made redundant in the same manner as those that employ detonating cord, so shock-tube systems must be connected together with particular care.

Detonating cord

Detonating cord contains a core load of high explosive (usually PETN). It detonates at about 22,000 ft. per second. Detonating cord is made with various weights of PETN per ft. of cord. This is usually expressed as the grains per ft. There are 7,000 grains in one pound.

Detonating cord is used as downlines in the blasthole to transfer initiation energy to primers and down-the-hole delays. It is also used for surface trunklines to connect blastholes together. It is easy to connect up, but has the disadvantage of generating substantial airblast. Therefore, it is usually used on surface when operating in remote locations. Shock tube, electric and electronic blasting systems are more commonly used when operating in proximity to built up areas.

Shock-tube systems

The shock-tube system is a plastic tube with a thin explosive coating on the inside of the tube. Upon detonation this material continuously detonates at a low velocity of approximately 6,500 fps. Thus, the plastic tubes are not consumed and the noise level is low. It is, therefore, good to use as lead-in line to initiate a non-electric blast in populated areas. It is also used to connect holes together when used as part of a long lead-surface delay system. It is used in the blasthole as a long lead down-the-hole (DTH) delay system to replace detonating cord downlines, or as a pigtail on DTH delays used in conjunction with detonating cord. Shock tube systems, unlike some detonating cords, will not set off a primer and must always be used with a DTH initiator and compatible primer.

Electric detonators

Fewer blasts in surface mines and quarries are initiated with electric systems today than once was the case. However, this practice is certainly still followed by many, especially in quarrying.

Construction of electrical caps and delays is similar to non-electric components, except that the energy to ignite the ignition compound is provided electrically. This does have the advantage of minimizing noise on surface, but has the disadvantage of being more susceptible to stray radio frequency and currents, lightening, etc.

The instantaneous electric blasting cap is sometimes called an E.B. cap. Like the non-electric blasting cap it is a thin metal shell containing various sensitive ignition powders and primary initiating high explosives sealed in a waterproof assembly. The electric cap is completely sealed with water-resistant plugs with only two insulated “leg wires” emerging. Inside the cap the leg wires are joined by a short piece of fine resistance wire called a “bridge wire.” The bridge wire may be imbedded either directly into an ignition mixture or in an electric match. In either case, when an ample electric current passes through this bridge wire it heats it to incandescence. This ignites the ignition mixture and initiates the primer and base charges in the cap. Thus, the electric blasting cap converts a relatively small amount of electrical energy into a primary-initiating explosion capable of detonating cap-sensitive high explosives with which it is in intimate contact.

Delay electric caps are similar to instantaneous caps in construction and action, except that between the ignition charge and the primer charge there is a column of powder called a “delay charge” which serves as a time fuse. Delay E.B. caps are of two general types: millisecond, and long-period delay. A wide choice of delay intervals are available running from about 8 milliseconds (a millisecond is one-thousandth of a second) through to about 12 seconds. Most quarries use millisecond delays because of the improved breakage and reduced vibration they provide. Many underground operations use the long-periods, although many have switched over to milliseconds.

Scores of different hook-ups may be made. Determination of electrical resistances and other details pertinent to firing electrically are discussed in manufacturers’ literature available to guide mine and quarry operators. Success requires that the operator precisely follow directions of the manufacturer who produced the electrical devices they utilize. Such directions give the exact procedure required to properly:

1. Select and lay out the blasting circuit.

2. Connect wires and protect splices.

3. Test the circuit.

4. Apply the required electrical energy.

5. Protect the circuit from extraneous electricity.

Electronic blasting systems

Both the shock-tube system and electric detonators rely on a pyrotechnic delay element to attain their delay timing. These pyrotechnic delays are subject to timing inaccuracies called “scatter.” Scatter can be caused by variations in the pyrotechnic composition, age and temperature. Deviation from the detonators nominal firing time can cause out-of-sequence firing. This will result in high vibrations, airblast and poor blast performance. Recognizing the accuracy issue and the safety concerns with both the electric system (stray current) and shock tube (cannot be tested) the industry has moved towards a more advanced initiation technology called electronic-blasting systems.

Electronic-blasting systems are unique as they have eliminated the pyrotechnic delay element and replaced it with a high-accuracy timing “chip.” These systems now deliver 1/10th of a millisecond timing accuracy with delays up to 20,000 milliseconds. The systems are available in both programmable and fixed times. Programmable systems allow the blast engineer to design blasts specific to the site conditions.

Electronic systems also bring with them many safety advantages such as being fully testable with self-diagnostics, able to operate in areas of extraneous current and greater blast control through accurate timing.

Field tests have proven that the use of electronic-blasting systems with proper blast designs have reduced vibration levels, airblasts and significantly improved blast performance.

Summary

The following summarizes the advantages of using delay detonators in production blasting.

• Improved fragmentation due to the greater freedom for the material to relieve.
• Greater flexibility in firing sequences and burden to spacing relationships due to the ability to orient the blast through the tie-in.
• Greater ability to control blast vibration and airblast.
• More predictable blast movement and flyrock control.
• Reduced backbreak behind the last row of holes.
• Minimized cut-offs.

Blast-design factors

There are a number of factors to be considered when designing a blast. These include:

• Material type to be blasted and the blast pattern and hole loading to use in the given rock.
• Degree of fragmentation desired.
• The geological structure and the attitude of the tie-in lines relative to the structure.
• The type and performance of the explosive charge.
• The type of initiation system and the duration of millisecond delays and accuracy needed.
• For a given pattern, the ratio of burden to spacing as defined by the tie-in or the layout.
• Subgrade drilling needed to fully break to the pit floor.
• Crest and toe locations (or average backbreak from the last row if the fact is not dug out).
• Upper bench elevations to determine hole depths.
• Blast size required to maintain quarry or mine production.
• Blasting ground vibration and airblast, and the design requirements to maintain acceptable levels.

Blasthole layout

Once a suitable pattern and loading have been determined it is important that the holes be accurately laid out in the field and drilled in the proper location. Irregular blasthole locations lead to less acceptable blasting results, unless the improperly drilled holes are redrilled. Burden and spacing dimensions vary and the tie-in is more difficult on irregular patterns. Some portions of the blast will be overshot, while other areas will experience hard toe and coarser fragmentation.

It is especially important that the front row holes be properly located. If there is too much burden (especially at the toe) fragmentation will suffer and the remainder of the blast will not properly relieve. Hard toes are likely to be in evidence. When there is too little burden, the high-pressure explosion gases cannot be contained. Rapid gas venting through the face will occur. Greater flyrock throw and airblast can be expected. There will be hard toes and blocky fragmentation.

Mines that have a survey department can measure toe and crest locations on the bench and plot these on the blast plan map, upon which drill pattern is designed. Thus, the front row locations can be more accurately determined. Quarries, where on-going surveying capability is less common can obtain a better idea of the face profile using a hand level and tape. Standing at the crest of the bench a point on the pit floor can be sighted and the angle measured with the hand level. Using simple trigonometry for the right triangle, the base length can be calculated since the bench height is known. This base length is the total horizontal distance from the crest of the bench to the point measured on the pit floor.

A 100-ft. tape is used to measure the distance from the point on the pit floor back to the toe of the bench. The difference between the total base length and this distance is the crest to toe offset. The blastholes can then be set back from the crest a suitable distance to yield an acceptable toe burden (or one can identify areas where an overburdened toe is likely to occur). In areas where safety working around the high wall is a concern the operation should consider laser profiling and bore tracking the blast to confirm the face conditions and profile. Laser equipment and or services are available to help acquire this information.

Blasthole loading

It is important that holes be loaded correctly in accordance with the design. Improperly loaded holes can lead to poor fragmentation and/or excessive flyrock and noise. The hole depths must be correct. Operators must decide how short holes can be before redrilling is required. In very hard rock a blasthole that is one or two ft. short can result in hard toe. In softer rock more variance is acceptable, but is seldom more than four or five ft.

Modern-day bulk trucks have more sophisticated measuring and control systems. The operator can set the weight to be loaded in the blasthole and the truck shuts off automatically. However, this does not eliminate the need to bob the blasting tape in the hole during loading. The truck-control systems cannot tell about voids or cavities in the hole, nor about control-system malfunctions. Thus to avoid over or under loading, and to obtain the correct column rise, it is still important to tape the hole during loading.

Accurate loading is especially important regarding the column rise and corresponding stemming height. The explosive column must rise high enough in the given rock type to fully break to the surface of the upper bench. Good breakage is related to the depth of burial of the top of the charge. Too great a depth of burial and the top of the blast will be poorly fragmented.

On the other hand, if the explosive column rises too high in the hole the depth of burial is low, gases vent rapidly to the surface, and there is more flyrock and noise. Also, the radius of the crater of fully broken material formed around the hole decreases and there may be hard areas between holes.

Front-row stemming height

Stemming heights on the front row may need to be increased. Since the bench-face angle is less than 90 degrees the burden on the front row holes is continuously decreasing between the toe and collar of the hole. Depending on where the front row blasthole must be drilled to maintain a suitable toe distance the burden may become too small to contain the explosion gases at the normal column rise. To avoid gas venting to the face causing flyrock, noise and loss of performance, stemming on front-row holes may need to be increased.

A simple measurement can be made in the field to determine the stemming height on the front row, using a telescoping surveyor’s rod and a 100-ft blasting tape. The rod is placed at the collar of the hole and extended over the face to the length of the desired minimum burden. A weighted tape is passed along the rod, through the ring at the top of the surveyor’s pole and then drops vertically until it strikes the face. The total length is read from the tape. The burden distance is subtracted from the total, giving the vertical distance. This is the stemming height. For example, if 6-1/2 in. holes at the quarry require a 12-ft. minimum burden and the total taped length is 29 ft., the vertical distance is: 29-12 = 17 ft.

This is the stemming height required to maintain minimum burden on the charge. Failure to make appropriate adjustments to front row stemming may well lead to hazardous flyrock.

Blast tie-ins and burden to spacing relationships

Drilled and loaded blast patterns may be tied-in to create different burden to spacing relationships. Commonly used designs are:

1. Square pattern tied en-echelon or across two free faces. Known as a V-1 tie-in. This is a non-staggered pattern. Tie-in is on the diagonal of the square and is oriented at 45 degrees to the free face. The effective burden is 0.707 times the drilled burden. The ratio of the effective spacing on the tie-in to the effective burden is 2:1.

2. Staggered square pattern tied on the diagonal of the parallelogram. This is known as the V-2 tie-in. The orientation is 34 degrees to the face. The effective burden is 0.56 times the drilled burden. The ratio of effective spacing to effective burden is 3:1.

3. Staggered equilateral pattern tied-in on the V-2 configuration. The angle to the free face is 30 degrees. The effective burden across the tie-in is 0.50 times the drilled spacing. The ratio of effective spacing to effective burden is 3-5:1.

4. Row on row tie-in. In this case successive rows detonate in progression. There is no burden reduction and the effective burden and spacing are the same as the drilled dimensions. The rows detonate parallel to the face rather than at an angle. Generally in open pits and quarries the tie-ins described above are preferred.

Millisecond delay timing

The duration of millisecond delay times must also be considered. Field experiments have shown that 1 to 1-1/2 ms-per-ft. of effective burden is the minimum that can be considered if any relief is to be obtained for holes firing on successive delay periods.

For good relief, it is typically found that 2 to 2-1/2 ms-per- ft. of effective burden are required. In some cases where maximum relief is desired 5 to 6 ms-per-ft. may be appropriate. When delay times are long care must be taken to avoid cutoffs and misfires depending upon the type of initiation system being used. A down-the-hole delay of sufficient duration to allow much of the surface tie-in lines and blasthole downlines to be consumed before blastholes begin detonating is the usual procedure taken to avoid these problems.

PREDICTION AND EVALUATION OF BLAST FRAGMENTATION

In recent years empirical methods for predicting the fragmentation that will be obtained from a given structural geology, rock type, explosive and blast pattern have become better and more useful. In addition, computer- assisted photographic techniques for measuring the size distribution of actual blasts have been developed.

It is good to recognize that there is currently no theory of fragmentation developed from first principles that can be used to accurately predict the fragmentation that will be obtained from a given blast. We seem to be at least a decade away from such a theory, and given all of the explosive, blast implementation and rock factors that influence fragmentation, one may well question whether such a theory is possible. For the time being at least the empirical approach is what we have to work with.

Empirical prediction of expected fragmentation is most often done using the Kuz-Ram approach. Using this approach one calculates a rock factor that describes the nature and geology of the rock. A uniformity index is also obtained that characterizes the explosive loading and blast pattern type and dimensions. This allows a characteristic size and size distribution to be calculated according to the Rosin-Rammler procedure. A considerable amount of fragmentation study is being performed in the industry using this procedure.

Actual measurements of fragmentation are often done today using the edge technique to analyze photographs or video of a blast to determine the actual size distribution. While these techniques may not always be totally accurate for the fines fraction they do provide a good overall assessment of the blast fragmentation. When photographs or video are taken at different times during the digging of the blasted rock a good idea of the degree of fragmentation throughout the blast can be obtained. Digital-fragmentation systems are now available to the operator that will allow them to monitor the fragmentation at the muck pile or at the processing plant.

A powerful approach appears to be combining predictive techniques with actual measurements of the same blast. If the two curves do not agree adjustments can be made to the rock factor in the Kuz-Ram calculation to bring them closer. Thus actual measurement can be used to “fine tune” the predictions. Then, when explosive, pattern size or other changes are contemplated better prediction of the resulting fragmentation will be achieved.

Fragmentation prediction and analysis have become increasingly important as it has been realized that primary breakage is less costly than secondary breakage, and often may be more economical than crushing. In addition, crusher throughput is directly related to the size distribution of the feed in relation to the crusher opening size.

The fragmentation distribution is very important to efficient run-of-mine heap leaching. When rip-rap is desired one may wish to look for combinations of geology, explosive and blast pattern that lead to blocky, coarse fragmentation. In quarry operations, excessive fines detract from the ability to produce products of more value and may create an environmental problem with regard to dust generation.

Authors: George D. Raitt, Robert McClure and Lyall Workman

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