For the purposes of this chapter, rock is defined as material which requires loosening by explosives or rippers in order to be dug economically by machinery.
Surface excavation of rock is done chiefly for the following purposes:
1. Stripping—the removal and wasting of any type of rock or dirt in order to uncover valuable layers.
2. Cutting—removal primarily to lower the surface. In road and airport construction the spoil is generally used for fill elsewhere on the project. In ditching, it is often used for backfill after installation of pipes.
3. Quarrying or mining—excavation of rock which has value in itself, either before or after processing. A rough distinction can be made between these two, in that quarries are ordinarily concerned with the physical characteristics of the stone, and mines with its chemical composition. However, the terms will be used interchangeably in this discussion.
One excavation can involve all three classifications, as in a heavy road cut where some material is wasted, other is used for road fill, and the best rock is crushed and used for aggregate.
Blasting may be divided into a primary operation in which rock is loosened from its original position in bulk, and secondary work which consists of reducing oversize fragments and breaking back ridges and spurs. The latter is done in the same manner as other light blasting, such as breaking boulders and chipping out ledges.
Rock work may also be classified as to the type and fineness of breakage required. Quarrying of building or dimension stone involves loosening large solid pieces from the parent rock, while blasting for fill or crushed rock requires pieces small enough to fit in the shovel bucket, the fill layer, or the crusher.
Stripping. In most stripping work the spoil has no value, so that the cheapest way of handling it is the best way. It is often possible to dump it in excavated areas from which pay material has been removed.
It is common practice to shoot and dig overburdens over 100 feet (30.5 m) deep in a single layer, and the use of the largest shovels and draglines is required for such work.
Drilling may be done horizontally from a face, as in Fig. 9.1(A), vertically from the top, as in (B), or in combination, as in (C).
Horizontal drilling has its best use when the mineral deposit is immediately under soft rock. Auger-type drills with extensions 6 to 10 feet (1.8 to 3.0 m) long, and diameters of about 5 inches (12.7 cm), are used. These have a tendency to drift downward, and since distances of 30 to 75 feet (9.1 to 22.9 m) are commonly drilled, it is necessary to start them several feet above the deposit, or to start them at a slight upward slant. Spacing may be 10 to 30 feet (3.0 to 9.1 m).
FIGURE 9.1 Horizontal and vertical drill holes.
For harder bottom rock, crawler-mounted (track) air drills may be used. The hole is usually started several feet above the bottom. Lower holes are slanted downward. Higher ones, if drilled, may be horizontal or slant upward.
If thick, higher layers of the overburden are hard rock such as sandstone or limestone, holes are drilled vertically, as in (B) or (C), using rotary or downhole drills with bits from 4 to 12 inches (10.2 to 30.5 cm) in diameter.
Burdens (distances from face) from 25 to 35 feet (7.6 to 10.7 m) and spacings in rows from 25 to 35 feet (7.6 to 10.7 m) are common in high faces. Track drills are used generally to depths of 50 feet (15.2 m), and occasionally to 100 (30.5 m). (See Fig. 9.2.)
Scrapers or dozers are sometimes used to remove most of the loose soil before drilling. This saves a considerable drilling footage, makes castings unnecessary, and, on low and medium faces, simplifies the use of track drills.
If the pit area is to be regraded when work is completed, the scrapers can be used to fill the hollows between the piles left by the shovels, or to place topsoil over regraded areas.
Because of the tremendous size and power of the excavators used in large pits, the blasting need only shake up, crack, and loosen the overburden without producing fragmentation comparable to that required in a cut or quarry. Wide-spaced holes and light charges can therefore be used successfully.
FIGURE 9.2 Drilling near a high face.
Highway Cuts. Rock cuts for highways may be of the through type as in Fig. 9.3(A), and the sidehill (B). Material from a sidehill cut may be thrown down to make a fill, as in (C).
The area to be cut should first be cleared and stripped of loose soil, and preferably of rotten rock. This may be done with dozers, scrapers, or shovels, depending on the conditions and the equipment available.
If the rock is soft, its upper surface may be loosened with rippers and removed, along with any dirt pockets it may contain. If it is hard and irregular, extensive cleaning by hand and with small equipment may be necessary. It is desirable to remove all loose dirt, particularly if the rock is to be used as crusher aggregate for road topping.
When water and disposal areas are available, hydraulic cleaning with contractors’ pumps and fire hose or with special equipment may be used.
If cleaning is not practical over the whole area, the spots to be drilled can be cleaned individually. The top layer may then be drilled, shot, and removed for fill, and any clean rock required can be obtained from lower levels.
If the cut is 20 feet (6.1 m) or less in depth, it may be taken in a single layer, but depths of 12 to 15 feet (3.7 to 4.6 m) are generally considered most satisfactory for track drills, and digging by 1- to 2½-yard (0.76 to 1.91 cu.m) shovels or medium-large front loaders.
In a through cut, the full width is used as a face to provide maximum space for machinery. On a sidehill, the same technique or one or more bench faces parallel with the centerline may be used.
Degree of fragmentation required is determined largely by the depth of fill layers where the spoil is used.
Mining and Quarrying. Pit operations are largely conducted to obtain certain classes of rock or earth. The general aspects of this work will be discussed in the next chapter.
FIGURE 9.3 Through and sidehill cuts.
Rock excavation may follow highway techniques in exploiting comparatively narrow or irregular veins; or large-scale stripping work may be necessary to make pay rock accessible to surface digging units.
Pits are often distinguished by the use of high and wide faces; or holes sunk below surrounding grades, with access by ramps or inclined or vertical hoists.
It is advantageous to have the face wide enough that several operations can be carried on in different sections with minimum interference.
The fineness of fragmentation which must be obtained by blasting is generally determined by the size of the hoppers or grizzlies on the crushers or processing machines.
The simplest type of drilling pattern is a straight line of vertical holes parallel to a vertical face. The distance from each hole to the face is called its burden, and the distance between holes their spacing.
The holes are drilled somewhat more deeply than the face so that any ridges left between them will not project above the new grade.
Blasts tend to overbreak at the top and not shatter completely at the base. As a result, faces tend to slope back. The projection of the bottom beyond the vertical line from the top is called the toe.
The extra burden at the toe may be handled by bottom drilling, or heavier loading (more powerful explosive or tighter packing) in the bottoms of vertical holes.
Holes may be drilled at an angle so that they are parallel to the slope of the face. This angle may be from 5 to 30 degrees.
Angle or slope drilling keeps the burden at the toe from being greater than at the top, so that no especially heavy charge is required at the bottom. Improved bottom breakage reduces the need for drilling below the floor, so that the total length of hole is usually not greater than in vertical drilling.
Cleaning. Small and medium air drills are not suited for working through soil overburden. It tends to choke up their air passages and to fall in on top of the bit. If this falling occurs after penetration into rock, it may make pulling the steel difficult or even impossible. Soil drills readily only when frozen.
The whole area is usually machine-stripped, and the actual spots to be drilled are cleaned by hand. At other times the soil is cleaned off only in slots, and drilling is done in lines along them.
Face Height. A rock mass may be taken out in a single layer, or in a series of benches. Highway contractors usually prefer benches, as do most open-pit miners, but quarry operators may take single slices 100 or even 200 feet (30.5 or even 61 m) high. This is partly because quarry rock is often sound enough to stand at great heights without much danger of collapse.
High faces are usually developed by pushing low or moderate ones back into a hillside, as in Fig. 9.4(A). Where low faces are preferred, such a cut may be made in a series of benches, as in (B) and (C). This is the safest method in most formations. The final slope should not be so steep that it cannot stand by itself.
Face height affects the method of drilling and the size and placement of holes. Accessibility of the top may be a determining factor in taking off the top layer.
In general, hand drills can be used efficiently for holes up to 10 feet (3.0 m) deep, and they can be carried anywhere a person can go. Their comparatively light air hoses can be strung over long distances, although at the cost of reduced pressure.
Track-mounted drills work well from 1 to 50 feet (0.3 to 15.2 m) deep, and may go to 100 (30.5 m) or more. They can climb and work on slopes over 30 degrees. However, risks and delays are involved in working them on very steep or rough ground, and it may be more economical to take a first cut with hand drills.
Depths over 30 feet (9.1 m) call for blast hole drills, of the rotary, downhole, or churn types. These heavy and expensive machines should be kept on safe, fairly even ground. They can dig through any depth of overburden readily, and can start their work from pioneer roads notched into soil slopes by dozers. See Fig. 9.5.
Hole Size. Hand drills will produce hole diameters of 1 to 2 inches (2.5 to 5.1 cm), wagon and track drills 1½ to 5 inches (3.8 to 12.7 cm), and blast hole drills 4 to 12 inches (10.2 to 30.5 cm). Pneumatic drills of all sizes make tapering holes with steel bits, as the hole gets smaller toward the bottom as they wear. Carbide bits and rotary and churn drills produce holes with little or no taper.
Borehole Loads. Figure 9.6(A) shows the cubic foot capacity of holes of various diameters for each linear foot. For example, to find out how much bulk of explosive will be needed to fill a hole 3 inches (7.6 cm) in diameter and 9 feet (2.1 m) deep, find the value 3 in the first column. Next to it is the capacity of 1 linear foot (0.3 m), .05 (
) cubic foot (0.0014 cu/m). Multiplying this last figure by the 9-foot (2.7-m) depth (below stemming) of the hole, we find that the capacity is 0.45 or 
cubic foot (0.013 cu.m).
The rest of the columns indicate the weight of explosive of various densities that it will take to fill the hole completely. The actual load will usually be somewhat less, except with free-running materials, because of waste space around cartridges. The amount of difference will depend on the efficiency of tamping.
FIGURE 9.4 High face and bench cuts.
FIGURE 9.5 Track and blast hole drills.
To find the amount of dynamite with a rated density of 164 sticks to the 50-pound (22.7-kg) case that can be tamped into this hole, follow the horizontal line in the table from the 3-inch (7.5-cm) line to the right to the fourth column, under the values 47 and 164. We find that a 3-inch (7.5-cm) borehole will hold 2.3 pounds (1.04 kg) of this powder per lineal foot (0.305 m). Multiplying this by the 9-foot (2.7-m) length, we have a possible load of 20.7 pounds (9.4 kg).
FIGURE 9.6 Borehole capacity and burden.
Untamped Cartridges. If cartridges are not slit and tamped, capacity is figured on the basis of the length and weight of the cartridges, and the number that can be placed side by side. Three of the standard 1¼ × 8 inch (3.2 × 20.3 cm) cartridges would fit in the hole if tied together. It would take 13 of these bundles to fill the hole within 4 inches (10.2 cm) of the stemming line. There would be 39 cartridges, each weighing 
of 50 pounds, or about 0.3 pound (0.14 kg) per stick. The total load by this method would be 11.7 pounds (5.3 kg), just a little more than half of the possible load.
Burden. The explosive in each hole is supposed to break out a section of the rock between the line of holes and the face. Only experience with the particular rock and explosive will indicate exactly the amount and type required.
In a general way, however, it may be said that 1 pound (0.45 kg) of 40 percent dynamite should break up and move 2 yards (1.53 cu.m) of soft rock, or 1 yard (0.76 cu.m) of medium hard rock, on an open face. In soft, layered, or rubbery rocks, 20 percent dynamite might move more per pound (kg); while in very hard rocks, even higher-strength dynamites might have smaller production. In tight holes, at edges, and in corners heavier loading is required.
Figure 9.6(B) is a table showing the yardage (cubic meters) of rock to be moved per foot (meter) of hole for various burdens and spacings.
According to the table, 3.33 yards of rock (8.36 cu.m per meter of borehole) would be moved if spacing were 9 feet (2.74 m) and burden 10 (3.05 m), or the other way around. An 8-foot (2.44-m) spacing and 12-foot (3.66-m) burden would yield 3.55 yards (8.91 cu.m).
Pull. If a hole fails to move its burden, it is said that it did not “pull.” This usually occurs at the bottom of the hole, and most often in edge or corner holes where the rock is held back on two or three sides. Such failures may be due to too heavy a burden or too wide a spacing, to improper stemming of a shallow hole, use of the wrong explosive, explosive not reaching the bottom of the hole, or a partial misfire. It is generally necessary to remove the blasted rock, check for any unexploded charge, then drill and shoot the bottom again.
Measurement. In order to drill and load holes accurately, it is necessary to know the height of the face and the amount of the toe. With low faces, or in casual operations, or in working upper lifts to temporary grades, depths may be estimated, although this is always risky.
Faces between 10 and 70 feet (3.0 and 21.3 m) may be measured by the device and method shown in Fig. 9.7. A 45° right triangle is carefully made of two-by-twos or two-by-fours, with the sides of the right angle equal and from 2 to 4 feet (0.61 to 1.22 m) long.
FIGURE 9.7 Measuring face height.
This is placed on the top edge of the face, as shown, and the bottom carefully leveled. A sight is taken along line A″AB and the spot B marked on the quarry floor by an assistant. The distance AB is then measured with a steel tape. Multiplying its length by .71 (the sine of 45°) gives the height of the face and the distance BC. The distance BD is then measured, and when subtracted from BC, gives CD, the projection of the toe.
This measurement may be repeated at various points along the face.
If the face is high, its top irregular, or considerable accuracy is necessary, it may be preferable to make a transit survey of the site, establish benchmarks and location points on all levels, and calculate from direct measurements from these points.
Each drill hole may be marked according to the cut to bottom grade, or by the drilling depth desired. For convenience in loading, the projection of the toe may also be noted on the marker.
Bottom Grade. The bottom grade should have a slope for drainage that may be away from the face or toward the sides, but not toward the face. If natural drainage is not possible, adequate pumps should be provided.
If blasting and excavating are accidentally carried below grade, hollows can be readily filled with fine shot rock. If the floor is too high, and the rock is soft, it may be possible to take it down with rippers and dozers. If it is too hard for machinery, a tedious job of shallow drilling and blasting, or of noisy and inefficient mudcappings, is required.
Spacing. In general, large drill holes are increasingly prevalent in faces over 30 feet (9.1 m) high, with proportionate increase in spacing and burden.
With most rock types, a point will be reached where enlarging and spreading the holes will result in poor fragmentation in the center of the blocks.
Faces 50 feet (15.2 m) or higher are usually shot with a single row of holes at a time. Low faces use additional rows, fired simultaneously or in short-interval succession. In any height, holes in the same row are now fired in sequence.
Each blast should supply enough rock to keep the shovel busy for at least half a day; therefore, the lower faces must be shot back more deeply than the high ones, particularly if the shovel is large and its working area narrow.
The entire group of holes may be drilled on a rectangular pattern, as in Fig. 9.8, or they may be staggered to improve fragmentation.
Best results from multiple rows are obtained if there is a free cleavage plane at the new grade. If the bottom is very hard to pull, heavier loading may be required than when firing single rows; or burdens may have to be reduced progressively toward the back, resulting in higher costs.
Computerized Drilling. Atlas Copco introduced its new system for surface drilling in Europe in 2007 and in the United States the next year. It is called the Rig Control System (RCS) with Global-Positioning (GPS) which delivers Hole Navigation for surface drilling. In the United States it is estimated that the combined system will save the $100,000 price tag in three years of use for a quarry.
The computerized system navigates the drilling equipment through a 10 ft × 14 ft (3.0 m × 4.3 m) surface pattern with extreme accuracy, and the holes are definitely vertical, or at a predetermined angle, regardless of finding an unexpected hard rock or a void in the drill’s path. The system can also be used for accuracy in drilling from a multi-drill jumbo equipment in a horizontal tunnel setup.
Tight Holes. When blasted rock must be sheared away on two or more planes, the shots are called “tight.” In Fig. 9.8 the holes marked a are open, those marked b are tighter, having to shear off the back or side as well as the bottom, and the c holes must shear along back, side, and bottom. In general, the tighter the hole, the greater the likelihood that it will fail to pull. It is usually cheaper to take special precautions with tight holes in the original blast, than to do secondary drilling and blasting.
FIGURE 9.8 Open and tight holes.
The tightest blasting found in open work is the start of a cut down from the surface: The first rock blasted can move upward only. If the whole set of drill holes are shot together, each of them will be very hard to pull. However, if the center holes are made oversize, loaded more heavily, and fired first in a delay sequence arrangement that progresses toward the outer limits of the area, the adjoining holes can throw into the space left, then the next holes throw into their space.
Buffers. Blasted rock may be entirely dug away before the next blast, or varying quantities may be left against the face. Complete cleanup is required if the toe is to be accurately measured or drilled horizontally, and is considered good practice with high faces.
If the working space is narrow, and the face low, it may be desirable to leave some shot rock as a buffer or “blanket.” This confines the force of the explosion, and may prevent blocking of the work area and aid fragmentation. However, the buffer should be small enough that some scattering occurs, as this makes it easier to find and blast oversize rock before the shovel gets in the heap. Heavier loading is required.
Snake Holing. Sometimes a face can be most economically blasted from the bottom only. This may be when a cliff (face) is being cut into a steep hill with poor access to the crest, the rock has a vertical cleavage so that it will break away without dangerous overhang when the bottom is blasted out, and is brittle enough to shatter when it falls. Such an operation may be very economical of powder, as the breaking out of the crest and the fragmentation upon hitting the floor are accomplished partly by gravity.
Coyote Holes. Coyote holes, illustrated in Fig. 9.9, are used for heavy undermining blasts. They may be used alone to topple a cliff, or to break out heavy toe burdens in conjunction with well drill holes loaded from the top. Faces should be at least 60 to 80 feet (18.3 to 24.4 m) high to justify their use.
Coyote holes are used when material is difficult to drill, a large yardage is required at one time, and the rock fractures readily.
They consist of small tunnels, 3 or 4 feet (0.91 or 1.22 m) in diameter, driven horizontally into the face at floor level, and one or two cross tunnels parallel to the face. Explosives are stacked in the cross tunnels, and the entrance is securely blocked with stemming. Firing is best done by Primacord.
Overbreak. Rock usually tends to overbreak at the top of the bank, and special drilling or loading may be required to avoid leaving a hard bottom rib at each blast junction.
It may be necessary to set back the first row of holes for the next blast for more than the normal burden, as their burden may be partly shattered, so that the shot is not confined as well as in the other holes. Drills may not be able to work close to the edge, or may not be able to penetrate shattered material.
In such conditions the front row may not pull the bottom unless it is drilled considerably deeper than the others, some horizontal bottom snake holes are drilled, or a denser or faster explosive is used.
However, shovels are frequently able to scrape away a few feet (meters) of unblasted soft rock, without excessive wear, in which case ribs may be ignored in the blasting pattern and dug out when found.
Edges. In large-area blasting, control of overbreak and underbreak is mostly a matter of getting the most efficient use from equipment and explosives.
In road cuts, and other work where remaining rock is a permanent part of the finished work, accurate finishing may be required by the contract, either directly as a specification, or indirectly in the form of nonpayment or penalties for excess removal.
In most rock formations, good to precise faces may be cut by line drilling, cushion blasting, preshearing, and various combination techniques. However, selection and refinement of method for a particular formation may be difficult.
For success, the rock must be reasonably sound and cohesive, and capable of standing at the slope to which it is trimmed. In starting a new project, it is highly advisable to obtain advice from your explosive company’s representative, and from local sources.
It is essential that the drills and the drillers be able to keep the holes in accurate alignment, both to obtain the desired edge and because the holes often show in the completed work.
Line Drilling. In line drilling, a single row of holes, usually 3-inch (7.6-cm) or less, is drilled along the edge, either from top to bottom, or in benches. Spacing is very close, from 2 to 4 hole diameters.
These holes are left empty. They create a line of weakness along which the rock should be sheared by the preliminary blast.
The back row of primary holes may be closer to the line than the normal between-row interval, and may be more closely spaced and lightly loaded than other rows.
This method is best suited to formations with a minimum of bedding planes, joints, and other weaknesses which may affect breakage more than the line of holes. Weak-bedded rock may respond satisfactorily if its planes are nearly at right angles with the proposed slope, or exactly parallel to it.
Cushion Blasting. Drilling for cushion blasting is similar to that for line drilling, except that larger holes and wider spacings may be used. (See Fig. 9.10.)
The holes are loaded with light charges, with undersize and/or spaced cartridges, often strung out on detonating cord. All space not occupied by explosive is usually filled with stemming, although some blasters prefer to stem only the top. See Fig. 9.11.
FIGURE 9.10 Preparing for a cushioned blast.
FIGURE 9.11 Charge placement for cushion blasting.
Closer drilling or heavier loading, or both, are needed in cutting angles in the face. (See Fig. 9.12.) Extra unloaded relief holes may be put between the loaded ones.
Standard procedure is to blast and usually to excavate the main cut back to within a few feet of the final line, either before or after drilling the cushion holes, and fire this final row separately. Detonating cord or instantaneous caps are normally used. If shock and noise are problems, closeinterval DuPont MS delay caps may be substituted.
Preshearing. Preshearing resembles cushion blasting in drilling and loading, but firing is in advance of the main blast. Since the explosive force has nowhere to go (except up), it may be expected to produce finer and more even fragmentation between the holes, and a totally effective shear plane for the primary blast. The two blasts may be fired together, using delay caps.
General Properties. Explosives are chemical compounds that can decompose quickly and violently. The original solid or liquid chemicals are largely changed into gases, including steam, that have a much greater volume. Heat is generated by the change, and serves to expand the gases greatly.
Explosion by rapid burning is called deflagration, and by almost instantaneous decomposition is called detonation. High explosives detonate.
Handling explosives requires experience. A widely used reference is Blaster’s Handbook, published by International Society of Explosives Engineers.
Properties to be considered in selecting an explosive include sensitivity, density, strength, velocity, water resistance, fumes, price, and availability.
Sensitivity is a measure of the ease with which a substance can be caused to explode and its capacity to maintain an explosion through the length of a borehole. It is also a measure of safety—the higher the sensitivity, the greater the risks of handling.
Nitroglycerin is so sensitive that it must be mixed with other substances before it can be used in commercial explosives. Compounds such as fulminate of mercury and lead azide that are used in detonators are so sensitive that they will explode at a light hammer blow or when exposed to moderate heat. At the other extreme, ammonium nitrate is so insensitive that few precautions and no permits are required for shipping and storing it.
Density is the volume of an explosive in proportion to its weight. It is measured in pounds per cubic foot (kg/cu.m), or in the number of 1¼ × 8 inch (3.2 × 20.3 cm) cartridges in a 50-pound (22.7-kg) case for wrapped dynamites. Such a cartridge contains about 9.78 cubic inches (0.96 liters). A count of 100 sticks to the case is roughly equivalent to a density of 77 pounds per cubic foot (1,230 kg/cu.m). A cubic foot (cu.m) of water weighs 62.4 pounds (998 kg).
Strength is the energy content of an explosive in relation to its weight. In general, maximum explosive power can be obtained from a given borehole by using a high-density, high-strength explosive in it.
Velocity is a measure, expressed in feet per second, of the speed at which the burning or the detonation wave travels through an explosive. It varies from 1,000 to 3,000 feet (305 to 914 m) per second for black powders to 23,000 feet (7,010 m) per second for blasting gelatin.
Low-velocity explosion has a heaving and separating effect, while high velocity crushes and shatters.
Water resistance is an important factor in wet rock, and varies with not only the character of the explosive but the manner in which it is packed and wrapped. Manufacturers are able to put water resistance in the explosive rather than in the wrapper.
The gases resulting from explosions vary in toxic and irritating qualities. This is very important in underground work, particularly if ventilation is poor. Explosives are rated by the manufacturers according to fumes as excellent, good, fair, and poor.
Explosives vary widely in the length of time they can be kept under various conditions before deterioration makes them dangerous or useless. Dynamite was formerly damaged by freezing, but this difficulty has been entirely overcome. Spoiling may be a serious factor if use is subject to delay, particularly in hot, wet weather.
Different explosives vary widely in price. The most economical one for a certain use may be one with a high cost. It is important to select an explosive on the basis of all the related factors, rather than purchase cost alone.
In many areas, very few types of explosive may be available, and because of the complications of shipping, delivery of special orders may be delayed weeks or months. Under such conditions, use of the standard dynamite may be advisable even if it is not exactly suited to the job.
If special explosives are purchased from a contractor or a quarry, it may be necessary to handle the transaction through a dealer in explosives to comply with state laws.
Permissible dynamites are those approved by the U.S. Bureau of Mines for use in gassy and dusty coal mines. Their most important feature is minimum flame in the explosion.
Black Powder. The explosives which explode by burning are called low explosives. Black powder is the only commercially important member of this class, and is the oldest explosive known.
Black powder is ordinarily composed of sodium nitrate, sulfur, and charcoal, finely ground and combined in grains of various sizes. The grains are then coated with graphite or other glazing to make the powder free-running. A more expensive powder for special purposes uses potassium nitrate instead of the sodium compound.
Fine-grain powders burn and explode faster than coarse-grain ones. Somewhat more powder can be packed in a borehole by mixing two or more grain sizes.
Black powder can also be obtained in pellets, which are short cylinders of compressed powder with a center hole. They are wrapped into 8-inch cartridges resembling dynamite in appearance. They are more convenient to use in small boreholes than the loose powder, and are somewhat safer to handle.
Black powder may be ignited or exploded by flame, heat, sparks, or concussion, and requires more careful handling than most dynamites. A special hazard is that powder spilled on the ground or on the magazine floor may ignite if stepped on or scuffed.
The blasting action will depend on the degree of confinement, the bulk, the grain size, and the closeness of packing. Unconfined powder will flash-burn, without explosion; and poorly confined powder will waste much of its energy along the path of least resistance.
Black powder produces considerable smoke and quite toxic fumes, the quantities of which vary considerably in different blasting procedures.
Black powder can be used to advantage when large, firm pieces of rock are desired, or when the material being blasted is soft and resilient enough to absorb the shattering blow of high explosives.
It cannot be used underground where ventilation is poor, or where the air may contain inflammable gas which may be ignited by the flame from the powder, or in wet holes. It has been replaced by other explosives in most applications.
Dynamite. Dynamite is the best known and one of the most widely used commercial high explosives. The name includes several different chemical groups, wrapped and marketed in about the same manner.
The “straight” dynamites consist primarily of a mixture of nitroglycerin, sodium nitrate, and combustible absorbents such as wood pulp, wrapped in strong paper to make a cylindrical cartridge. Although a wide variety of sizes are available, the most popular are 8 inches (20.3 cm) long and 1⅛ or 1¼ inches (2.86 or 3.18 cm) in diameter.
The percentage of nitroglycerin by weight contained in the mixture is used to identify it, according to strength. From 15 to 60 percent may be used.
Strength does not increase in proportion to the percentage of nitroglycerin because the other ingredients also contribute gas and heat. For example, a 60 percent dynamite is about 1
times as strong as a 20 percent.
Higher percentages are faster and more sensitive. Speed is desirable in hard rock and where the explosive is not confined, as in mud-capping boulders. Sensitivity is necessary when blasting mud ditches by the propagation method.
Straight dynamites have fair water resistance. Their fumes are poor, however, and they are never recommended for underground work.
Any type of dynamite of the general-purpose 40 percent strength will explode if subjected to sharp concussion, such as explosion of a blasting cap; from impact of a rifle bullet; from excessive heat, whether produced by fire, friction, or impact; and from sparks.
When dynamite is burned—usually to destroy surplus or deteriorated stock—it is spread in a thin layer on straw or other combustible material, which is ignited. All personnel should keep a safe distance from the fire. Dynamite will usually burn without incident, but there is always a chance that it may explode.
Spoiled dynamite may soak into its containers, and render them explosive. The cases and wrappings should therefore be burned with the same precautions as would be taken with dynamite.
The “ammonia” dynamites use ammonium nitrate as the principal explosive, in combination with some nitroglycerin. They do not catch fire as easily as straight dynamites, and are less sensitive to shock and friction. Water resistance is generally inferior, but fumes are less objectionable.
These are rated on a percentage strength basis, but the figures do not indicate anything of their chemical composition but simply that performance is comparable to that of a straight dynamite of the same rating.
A third type of explosive used in commercial blasting is gelatin dynamite. This is based upon a jelly made by dissolving nitrocotton in nitroglycerin. Various other ingredients are added.
The gelatin dynamites are dense, plastic, cohesive, and practically waterproof. Fumes are excellent in all but the highest strengths, which vary up to 90 percent.
Shot with a standard cap, and when not confined, ordinary gelatin dynamites will explode at a velocity of about 5,000 feet (1,524 m) per second. If confined, or shot with a straight dynamite primer, velocities of 13,000 to 22,000 (3,960 to 6,700), depending on the strength, will be obtained. Certain types may also be obtained which will always detonate at the higher velocity.
Gelatin dynamites are relatively insensitive to shock, and often will not explode by propagation from adjacent holes. Their plasticity makes it easy to load them solidly in boreholes, and to pack tightly in cracks for mudcapping. The velocity of the higher grades and their high density recommend them for hard, tight blasting; and the waterproof qualities for any underwater work not requiring propagation.
Ditching, stumping, and agricultural dynamites are usually one of the standard strengths best suited for the purpose, with a special designation.
Blasts are initiated or set off by timing devices, by remote electrical controls, or by a combination of these methods.
Ammonium nitrate (NH4NO3), called AN for short, is a nitrogen fertilizer that has largely replaced dynamite in medium and large borehole blasting.
At ordinary temperatures AN is generally a stable compound. If heated to 300 to 400°F (149 to 204°C), it will decompose without exploding into water and nitrous oxide, a brownish red gas with a pungent smell.
If subjected to great heat under confinement, or direct detonation of high explosives, AN decomposes explosively into water, nitrogen, and oxygen. The formula is
2NH4NO3 = 2N2 + 4H2O + O2
Ammonium nitrate is so insensitive that it is not rated as an explosive. It is called a “blasting agent,” and it can be shipped and stored free of special regulations and permits. However, it does burn, and since it supplies its own oxygen, it cannot be smothered. The only way to put it out is to use plenty of water.
If burning AN is confined in rooms, or if it is piled in such bulk (over 100 tons to a pile) that heat and pressure can build up inside it, it may explode as destructively as dynamite.
AN should never be transported or stored with high explosives.
Additives. The sensitivity and explosive power of AN are greatly increased by mixing or blending it with organic materials that absorb oxygen when they burn. Several disastrous explosions have occurred in AN fertilizer that was treated with a small amount of wax to prevent sticking.
For use as an explosive, additives such as lampblack, powdered coal, sawdust, and fuel oil have been tried. Fuel oil is very successful, of which more will be said below. Any of these provide material to combine with the surplus oxygen to produce additional heat and gas volume.
Prills. Most AN is prepared in prill form. Prills, sometimes miscalled “pellets,” are globular, porous particles obtained by spraying a 95 percent solution of AN into a rising current of warm, dry air. They are usually coated with about 3 percent by weight of kieselguhr (diatomaceous earth, about 80 percent finely divided silica, called guhr for short) to prevent sticking. There are also uncoated prills, and prills coated with minute amounts of other materials. Prills vary in density, size, and size mixtures according to brand and specification.
Other processes produce flakes, grains, or pellets that usually are denser. These have not been as successful in direct blasting as the prills up until now.
History. AN has been used as an ingredient in dynamite and other explosives since 1867. In 1935 a canned AN mixed with a small quantity of fuel was introduced by Du Pont under the name Nitramon. Because of its low sensitivity it could not be detonated by caps, and special primers of amatol, a TNT and AN compound, were provided.
Nitramon and similar products are safe and clean to handle, economical, excellent in wet holes and coyote holes, but are now declining in use because of competition from fertilizer grade AN.
Nitrex, a canned AN, was used in a 3,300,000-pound (1,500,000-kg) blast to remove Ripple Rock above Vancouver.
Fuel Oil. In 1955 it was discovered that No. 2 fuel oil is an almost ideal material to mix with AN to make a practical explosive. The mixture is called AN-FO for convenience.
Maximum explosive power is obtained when the mixture contains about 94 percent AN and 6 percent fuel oil. This is about 1 gallon of oil to a 100-pound (45.4-kg) bag of fertilizer. This is also the proportion of oil that the AN absorbs most readily.
Mixing may be done at the factory, at a fixed plant near the job, by mobile equipment or by hand at the borehole, or in the borehole. Relative cost depends largely on the size of the job.
Factory mix is the most thorough and the most expensive, and is the least trouble on the job. The mixed AN-FO, called nitrocarbonitrate by the ICC, is more dangerous to handle than the unmixed AN. Various mild precautions must be taken, and vehicles carrying it must be marked “DANGEROUS.”
AN-FO should not be handled or transported in the original AN bags unless the change in contents is plainly marked.
AN-FO may be made dangerously sensitive, or even caused to explode spontaneously, after long standing in stemmed boreholes, by contamination with unidentified, naturally occurring chemicals.
Local plants should mix as well as the factory, and mobile equipment almost as well. Hand mixing is much less effective, and mixing in the borehole is rather hit-or-miss. A thorough blending is important for highest blasting efficiency, but it may cost more than it is worth.
Poor mix, or poor detonation for other reasons, is likely to be indicated by “Kodachrome clouds” of yellow, orange, or brown smoke from the explosion. The color is caused by nitrous oxide mixing with steam and other gases.
Borehole mixing is done by pouring in a bag or two of AN, adding the correct quantity of fuel oil, then putting in more AN and more oil alternately until the full charge is placed.
Hand mixing outside the hole may be done by pouring 3 quarts (2.8 liters) to a gallon of fuel oil into each 80-pound (30.3-kg) bag of the fertilizer, or 2 to 2½ quarts (1.89 to 2.37 liters) into each 50-pound (22.7-kg) bag, moving the bag around to help the liquid to distribute itself evenly, and allowing it to stand for 5 to 20 minutes before pouring it into the hole.
Somewhat better results may be obtained by putting the AN in a mixing trough such as is used to hand-mix concrete or plaster, spraying the fuel oil over it, and mixing with a hoe or shovel.
Unopened bags may have fuel oil injected into them under high pressure by the use of an engine-driven pump and a sharp-pointed nozzle that can be pushed through the wrappings. This method, pioneered by Monsanto Chemical Company, is called the Needle Fuel Injector System.
Mechanical mixing may be done with engine-driven cement mixers. Oil should be added as a spray rather than a solid stream.
A great deal of specialized equipment has been designed, both in the factory and in the field, for mixing quantities of prills with fuel oil, and discharging them rapidly into boreholes. Augers may move prills from a truck body into a stream of compressed air, in which they are mixed with oil and blown into the borehole through flexible hose.
Loading prills by compressed air may build up charges of static electricity in the borehole and its vicinity. Such charges create danger of premature discharge of electric caps, and even possible hazards with fuse caps. Minimum precautions are grounding of the pneumatic loader and use of semiconducting hose. Caps should be of types least sensitive to stray currents.
This problem, and a number of important precautions, are discussed in a U.S. Bureau of Mines Report, IC 8179, titled “Safety Recommendations for Sensitized Ammonium Nitrate Blasting Agents.”
Priming. Most AN-FO cannot be detonated dependably by blasting caps or regular Primacord. Even when detonated, it may not maintain full speed of explosion, or even any explosion, for the length of the hole.
It is therefore necessary to use primers or boosters that can be exploded by caps or Primacord, and that will produce sufficient explosion to detonate the AN-FO at high velocity; and to use enough of them in the hole that the explosion wave will not have enough space between them to weaken.
Such primers may be made of one or more sticks of gelatin dynamite of 60 percent or higher strengths, or of special “cast” explosives. The cast boosters are somewhat more expensive than dynamite, but are safer to handle and are more powerful.
Figure 9.13 shows a cross section of a cast booster. The sensitive explosive that is detonated by Primacord or a cap is completely surrounded and protected by another high explosive that cannot be detonated by a cap or any probable accident, but can be detonated by the core, and has sufficient strength to set off any AN-FO mixture.
Borehole diameter is the most important single factor affecting propagation of the explosion, as this dies out much more quickly in small holes than in large ones. Although AN-FO has been successfully used in holes as small as 2½ inches (6.4 cm), variable results, and cost of extra boosters, limit its use in holes smaller than 4 inches (10.2 cm).
Specification. AN-FO is a passive explosive material. Its performance is governed by such factors as confinement, diameter, and fuel oil content. Under defined conditions of density, confinement, diameter, and particle size distribution. AN-FO will have only one velocity at which it can be detonated. The maximum velocity at which it can detonate is 14,600 feet (4,450 m) per second in a borehole 10 inches (25.4 cm) or larger in diameter. In a 4-inch (10.2-cm) borehole, the steady-state velocity of AN-FO is 12,500 feet (3,810 m) per second, and in a 6-inch (15.2-cm) borehole it is 13,900 feet (4,240 m) per second. If the primer has a lower velocity, it may not be able to detonate the AN-FO. Ideally, the primer will have a higher velocity than the AN-FO, and there will be a downward change in velocity known as overdrive. Overdrive velocities have been recorded only in boreholes up to 6-inch (15.2-cm) diameter.
The other key factor in detonation of AN-FO is the diameter of the primer charge relative to the AN-FO explosive. As the diameter of the primer is reduced below that of the explosive, the effect on the initial velocity of the AN-FO is considerable. An extreme low velocity is developed in the AN-FO when a 1-inch-diameter primer is used. This suggests using a slurry-type primer that will fill the borehole as the AN-FO does. If that is not practical to overcome the effect of diameter mismatch, it is necessary that primers have a detonation pressure of at least 80 kilobars (more than 1,000,000 pounds per square inch). Primers satisfying that requirement include Aquaram, Aquanal, Power Primer, Pentolite, gelatin dynamite, and perhaps TNT or water gel, but not regular dynamite.
To summarize: The efficient initiation of AN-FO requires a primer that has sufficient diameter to nearly fill the borehole and has a minimum length of at least one borehole diameter; the detonation pressure of the primer should be at least 80 kilobars.
Specifics. The Du Pont company originally developed a family of high-explosive primers known as Detaprime, which stress safety, high performance, economy, and ease of handling. These primers have a velocity of 24,000 feet (7,320 m) per second and a density of 1.50. They are insensitive to mechanical impact, have no metal-to-metal contact between the primer and cap, and do not detonate when exposed to open flame. There is no sensitivity to temperatures from –40 to 225°F (–40 to 107°C). Their water resistance is excellent, including resistance to moisture throughout shelf life. And storage life is 3 to 5 years at ambient temperature. The Detaprime primers are about ½ inch (1.3 cm) in diameter and 2 to 3 inches (5.1 to 7.6 cm) long, with a hollow core for detonating cord, fuse, or cap insertion.
Poor confinement may also cause the detonation to slow or stop. A soft rock or mud seam between layers of hard rock may not confine the AN-FO sufficiently for it to carry the explosion across.
In general there should be a primer at the bottom and at the top of the hole, and one at least every 20 feet (6.1 m). However, 50-grain Primacord and a single primer at the bottom may prove adequate.
Loading. Dry holes are loaded by placing the primers with the detonating cord or wire, then pouring the mixture or the two ingredients separately until the proper amount has been placed. Stemming is then added, as described later under LOADING.
There is a tendency to load holes higher with AN-FO than with dynamite because of its lower cost, and to use less stemming. Overdoing this is likely to be a waste of the explosive, and to add to the hazard of high flying rocks.
Unprotected AN-FO cannot be expected to perform dependably in wet holes. It can be protected by putting it in plastic bags. It may be bought this way, or the bags made up from sheet plastic with the help of a sealing machine.
Bagged AN-FO tends to float on any water in the hole, and must be forced down by the weight of explosive above it. Even with firm tamping the bags will reduce the amount that can be placed because they will not conform perfectly to the walls. Bags may tear and allow water to ruin their contents, perhaps cutting off part of the blast.
Density. The density of AN ranges from 47 pounds per cubic foot (753 kg/cu.m) for the prilled variety to 64 pounds (1,026 kg/cu.m) for fine-grained types. This compares with dynamite densities of 37 to 90 pounds (593 to 1,448 kg).
In loading calculations the low density of prilled AN-FO compared to heavier dynamites is offset to a variable degree by complete filling of all hole space by the free-running material. That is, if a dynamite with a 60-pound (962-kg) density filled only 80 percent of the bore space in spite of slitting and tamping the cartridges, then its effective density in the hole would be .8 × 60 or 48 pounds per cubic foot (769 kg/cu.m), about the same as the prills.
The finer and denser types of ammonium nitrate are not well suited for use as blasting agents with present techniques. They are more sensitive and less powerful than the prills, and are more difficult to mix with fuel oil. Some increase in density of prills may be obtained by mixing two or more sizes, but the heavier charge may still be offset by slower detonation.
Composition. Slurries, also known as water gels or dense blasting agents (DBAs), are usually mixtures of a “sensitizer,” an oxidizer, water, and thickener.
The sensitizer may be any of a number of reducing (oxygen-hungry) chemicals. This is usually the explosive TNT (trinitrotoluene), but may be (or include) finely divided aluminum and/or other substances that may or may not be explosives themselves. They may be in sand-size granules, very fine powder, or other forms.
The oxidizer is an oxygen-rich chemical such as ammonium nitrate and/or sodium nitrate.
Characteristics. Consistency is regulated by the amount of thickener or jelling agent (often guar gum) that is used. It varies from that of pancake syrup to soft (flows if jiggled) jelly at room temperature. Stiffening occurs at low temperatures, but most formulas are resistant to damage by freezing.
Water resistance varies from good to excellent, unless water flow is sufficient to wash it away. Loading water gel in its sealed plastic packages is recommended for severe water conditions.
Packaged slurries are generally jelled in cylindrical shapes slightly smaller in diameter than the boreholes in which they are to be used. They are shipped in polyethylene bags, protected by burlap or by cardboard containers.
The bags are usually soft enough to allow slump to fill the borehole almost completely, and can be obtained in even softer, expandable types. However, it is normal to slit the top 6 inches of each bag, and drop it into the hole. Some water gels have unique gellant systems which are designed to permit them to be poured into the hole.
It does not mix readily with water, and its weight (specific gravity is 1.1 to 1.5) causes it to flow to the bottom, displacing any water and filling all borehole space. But if it is very thick, it may bridge over and retain water pockets.
For large-scale use, slurries may be shipped and even mixed in tank trucks and pumped into holes. If the hose can be extended to the bottom of the hole, danger of bridging is eliminated. Most slurries are insensitive. They cannot be detonated by blasting caps or Primacord, and require special booster-primers.
Recently developed small- and intermediate-diameter grades of water gel are cap-sensitive, so that they require shipment and storage as a class A explosive. This type is finding wide acceptance as a replacement for dynamite in bottom loads.
Applications. Slurries are less expensive than dynamite, but cost more than ammonium nitrate–fuel oil mixtures (AN-FO). They are now used chiefly in open-pit mines where rock is hard and/or holes are wet.
The high density of slurry (1.37 to 1.68) compared with AN-FO permits use of smaller-diameter boreholes, or wider spacing, to obtain the same explosive power and fragmentation. The higher price of the slurry may be more than offset by reduced drilling cost.
In wet holes AN-FO is not practical, so the choice is between dynamites and slurries. Here the slurries have a smaller advantage in density, but have a price advantage as well.
Some slurries have been employed successfully in boreholes down to 1½-inch (3.8 cm) diameter, using a special small-diameter booster.
Safety Fuse. The original timing device is a fuse made up of a black powder core, surrounded by layers of protective wrappings. Two speed ranges are available, with burning speeds of 90 seconds and 120 seconds to the yard. These speeds must be considered approximate, as they are affected by altitude, weather, storage conditions, and possible damage to the fuse.
Fuse is water-resistant except at the cut ends, which are immediately spoiled by contact with moisture. It should not be used unless it can be shot the same day it is loaded.
It is manufactured in lengths of 50 feet (15.2 m) or more, and wound in coils or on spools to be cut to the desired length on the job. As short an interval as possible should be allowed between cutting and using.
Squibs. Electric squibs are devices for igniting charges of black powder, and may be used instead of fuses. They consist of copper or aluminum tubes with powder, an electric firing element, and wires sealed into them. They are embedded in the powder charges, and when sufficient current passes through them, they catch fire and ignite the charge.
Blasting Caps. Figure 9.14 shows construction of blasting caps for both fuse and electric firing. Fuses are inserted in the hollow shell of the cap, and fastened in by crimping the metal with special tools or machines. The fuse should be cut off square, preferably by a razor blade or other very sharp edge, which will not pinch the wrappings together.
The electric firing device consists of a very thin wire lying in a highly combustible mixture. Passage of electricity causes the wire to become white hot and ignite the mixture, which explodes the primer and high explosive.
Delay Blasting. This is used to improve rock fragmentation and rock movement. Rock fragmentation is dependent on the stress wave set up in the rock by the explosion. That may be between 10,000 and 20,000 feet (3,050 and 6,100 m) per second—faster in harder rock. The rock fractures radiating out from the borehole travels from 0.15 to 0.4 times the stress wave in the rock, so they travel from 1,500 to 8,000 feet (457 to 2,438 m) per second. In other words, the radial cracks travel from 1½ to 8 feet (0.46 to 2.44 m) per millisecond.
As a guide to when rock movement starts, the common expression of 1 millisecond per foot (3.3 millisecond per meter) of burden is reasonably correct. The rate of movement is used as a guide so that rock broken by the first row of holes will move out of the way before the second holes fire, or material from the second holes has no place to go. The delay firing may be obtained by use of delay electric blasting caps. Time interval averages about 1 second in “standard” caps. A more effective delay system will be obtained by using MS connectors with 5- to 25-millisecond delays. The very short intervals give the finest breakage but give reduced efficiency and increase throw, making mucking of the round difficult due to widely scattered rock.
Wires may be obtained in almost any desired length, and should be long enough to reach the wires from adjacent charges or to connect with the lead wire to the electrical source. They may be copper or iron, and are protected by a plastic insulation. The ends of the wires are fastened together into a bridge or shunt, to prevent premature firing through contact with stray electric currents.
Instantaneous and delay caps may be used on the same round. If they are in the same series, it is good to increase firing current by one-third. Caps made by different manufacturers should never be fired together. They are almost certain to have different current requirements, so that one brand of cap will fire and break the circuit before the bridge wires in the other brand are heated enough to fire.
Caps contain explosives which are more sensitive than dynamite, and they must be handled with great care. Heat, friction, and shock must be avoided. In the original package, electric caps are usually cushioned in their own folded wires, and are often protected by a cardboard or paper wrapper in addition.
But the greatest danger of accidental firing of electric caps comes from electricity. Even with their wires shorted by a soldered shunt in the original package, a very powerful nearby current might detonate them unexpectedly.
After the shunt is opened to connect to other wires, there is danger from any stray electric current, even from radios.
The content of a cap is small, but one can blow off fingers and toes, and flying particles of the copper case may cause injury to personnel within a radius of 30 feet (9.1 m). The most serious danger in an accidental explosion of caps, however, is that of setting off primers or nearby heavy explosives.
Caps should be buried before exploding for test purposes.
Primacord. Primacord is a detonating (exploding) fuse, made up of a core of an insensitive high explosive, pentaerythritol-tetranitrate, that is called PETN for short, surrounded by a protective wrapping. Primacord is detonated by means of a blasting cap. The explosion travels along it at a rate of about 21,000 feet (6,400 m) per second, and detonates any cap-sensitive explosives with which it is in contact. See Fig. 9.15.
It is produced in a number of types that are classified according to explosive content in grains per foot, and/or the type of protective covering around the explosive core.
Standard types usually have 50 to 60 grains per foot (164 to 197 grains per meter). Lighter grades, including “E-Cord” with 25 grains, are used chiefly for secondary and very shallow blasting. Heavier ones, with 100 to 400 grains, are used for continuous column initiation of ammonium nitrate–fuel oil mixtures, or for cutting into short lengths for use as primers.
FIGURE 9.15 Primacord is supplied in boxes and rolls.
Wrappings are rated on the basis of strength and water resistance. All are water-resistant, but plastic-coated ones are essentially waterproof except at cut ends. The plastic is also resistant to oil, an important point when using an AN-FO mixture.
E-Cord and Primacord with 45 or more grains will fire when wet with water or oil if it is initiated from a dry spot. Wet Primacord can be initiated only with a very powerful primer, such as 80 percent gelatin dynamite or a special booster. It will not maintain detonation through a knot connection.
Cut ends of Primacord will pick up some moisture from capillary attraction, usually only a few inches in. If the cord is lower than the water level, it may in time become soaked all the way through.
Textile-reinforced Primacords are usually used in ordinary downholes, wire reinforcement in rough and jagged ones.
The individual lines in the holes, called branch or down lines, are usually connected to the blasting cap by a trunk line of Primacord. Fastening is done by simple knot connections. It is important that the lines be at right angles to each other. (See Fig. 9.16.) When very stiff cord is used, it may be necessary or convenient to use plastic connectors instead of knots.
Electric caps may be used to combine quiet surface wiring with Primacord downlines. Connection is made by taping a cap (or preferably two caps) tightly to the Primacord. This may be a short piece, called a tail, which is tied to the down line with a square knot just before firing. The caps are wired into a conventional blasting circuit.
There are also nonelectric delay caps, Primadets, that are made up attached to a light detonating cord, Primaline. They are strong enough to explode AN-FO and most slurries directly, without a primer.
The cap is placed near the bottom of the hole, with or without a primer, before pouring in the main charge of explosive. The Primaline is knotted to the Primacord trunk line. It does not have sufficient strength to detonate AN-FO or slurry, so timing is regulated by the delay feature in the cap. It is not used with dynamite or other explosives that are sensitive enough to be set off by the cord.
On all large blasts it is customary to arrange Primacord so that each hole may be reached by the explosion from two directions. Large blast holes, or holes of any size with deck loading, may have two cords strung on opposite sides of the hole to ensure firing.
Primacord has become the standard method of setting off large blasts, because of its exceptional safety. As an explosive it is quite inert, and is less likely to be detonated by accident than the main charge of explosives. Particularly, it cannot be exploded by stray electric currents, a serious hazard for electrical hookups. An entire blast can be prepared with detonating cord, and a fuse or electric cap attached at the last minute.
FIGURE 9.16 Primacord connections. (A) Trunkline spliced with square knot. (B) Double wrap, half-hitch knot connection between a reinforced trunkline and downline. (C) Trunkline clove hitched over a downline. (D) Special clove hitched trunkline to prevent slippage from subsidence of the charge.
Low-Energy Detonating Cord. The noise made by the explosion of Primacord trunk lines aboveground may be objectionable. This noise can be considerably reduced by covering it to depths from a few inches up to a foot with sand, dirt, drill cuttings, or other stone-free material, or almost eliminated by using low-energy cord.
LEDC, a low-energy cord that is relatively noisefree, was developed jointly by Ensign Bickford Company and E. I. Du Pont de Nemours Co. It contains 2 grains of PETN per foot (0.3 m), protected in a tiny lead tube covered by wrapping of cotton and plastic. The noise made by exploding 150 feet (45.7 m) of it is equal to that caused by one blasting cap or 2 inches (5.1 cm) of reinforced Primacord.
Special connectors are available to fasten LEDC to Primacord and to itself. When the explosion travels from LEDC to LEDC or to Primacord, a booster charge is needed. This may include a 0-, 10-, 15-, or 25-millisecond delay unit.
The preferred method of firing is to fasten an electric blasting cap to one of the down lines of Primacord, and to fasten the LEDC to this down line with a nonexplosive connector called a trunk line adapter. The LEDC is connected to other down lines by booster delay caps.
Primers. A primer may be a stick of dynamite that contains a blasting cap; or is any other heavy explosive which has been fitted with a device for setting it off.
Since these primers combine the power of the dynamite with most of the sensitivity of the cap, they must be handled with greater care than any other units of explosives.
They are ordinarily prepared at the borehole immediately before being placed, but may be made in some central place and delivered to the loaders as required.
The essentials of a good primer are that the cap must be powerful enough to produce detonation, there must be intimate contact between cap and explosive, they must be fastened together so that they will not separate while being placed, the cap should be shielded from shock or friction, and the wires or fuse should not be kinked or strained.
Black Powder. Black powder may be primed by placing a fuse in the hole and pouring the powder around it. This method may be improved by tying a knot in the end of the fuse to anchor it, and making several slits into the core above the knot where they will be in contact with the powder. A paper cartridge may be prepared to hold powder closely around the knot and slits.
If the powder is to be exploded by an electric squib, a similar cartridge is made up to enclose the squib with some powder.
Blasting caps may also be used to explode black powder.
Fuse Caps. Preparation of dynamite and fuse cap primers includes two jobs—attaching the fuse to the cap and the cap to the powder. The fuse ends must be dry.
The fuse should be cut squarely with a clean, sharp blade, preferably a razor blade in a suitable mounting, and pushed into the cap until it is seated against the explosive compound. The copper shell is then crimped firmly onto the fuse with a hand or a bench crimper.
If the fuse is cut on a bevel, it may fail to make proper contact, Fig. 9.17, top, or the end of the casing may curl over and prevent contact. A good contact is shown at the bottom.
The interior of the cap should be clean. Any foreign matter in it should be tapped out or removed with a straw or toothpick. Blowing into it may dampen it and cause it to fail. If the cap is suspected of being damp from any cause, it should not be used.
FIGURE 9.17 Insertion of fuse in cap.
FIGURE 9.18 Priming dynamite with fuse cap.
Figure 9.18 shows two ways to place and fasten a fuse cap in dynamite. In each case a hole or slit is made in the cartridge and the cap inserted. The primer can be held together by lacing the fuse through another hole, or by tying it with string.
In shallow holes, and in blockholing or mudcapping, it is practical to simply insert the cap in the cartridge end, without fastening, as the primer need not go out of reach. Friction will hold it in place against a moderate pull, but not against yanking.
Electric Cap Primer. Figure 9.19 illustrates the most common method of priming a small-diameter stick of dynamite. The cap is pushed into the navel-like end of the wrapper, and the wires are caught in a half hitch around the center. If the dynamite is hard, a hole may be made in the end with a wood peg to make it easier to insert the cap securely so that it will not slip out during loading.
Large-diameter sticks are best primed by this sequence: the cap is inserted in the top, and a loop of wire is pushed through a slanting hole and is caught around the middle.
In each case the cap is entirely inside the dynamite, cannot work into a position where it might scrape the side of the borehole, and its direction of explosion, away from the wires, is directed into the dynamite.
Detonating Cord. To make up a primer, Primacord is fastened to a large stick of dynamite by being threaded through and secured with tape and to a small one by being tied tightly. The resulting primer is usually the first explosive to be put in the hole. This puts the detonating cord in contact with the entire charge. See Fig. 9.20.
FIGURE 9.19 Priming small dynamite stick.
FIGURE 9.20 Primacord and a large dynamite stick. (Courtesy of Ensign-Bickford Co.)
Placing Primers. If one primer is to be used in a borehole, it is best to place it at the top, or one cartridge down from the top. This keeps to a minimum the danger of damage to fuse or wires while placing and tamping the charge. Use of one stick above the primed one cushions the primer against jars from overzealous tamping, and from contact with abrasive stemming material.
If the hole is long, or the charge heavy, it may be considered a good precaution to use two or more primers. The second one is liable to be in the bottom, and any additional ones are spaced throughout the column.
With all types of delay firing the primer must be in or near the bottom. This largely avoids the danger of throwing the primer out in the muck pile, where it would be likely to detonate during digging, with a possibility of disastrous results.
Spaced charges are more likely to require additional priming than solid ones. Deck charges need a primer in each level.
Correct placing of primers, and even correct direction of the cap in the primer, is of importance under some circumstances and makes little difference in others. Because of the speed and destructiveness of blasts, exact analysis of the mechanism and results is difficult.
Speed of Explosion. There are four classes of speed concerned in the firing of explosives. There is the slow burning of a fuse, explosive burning of confined black powder, extremely rapid but somewhat variable detonation of high explosives, and almost instantaneous passage of electricity.
Black powder is little used, but still serves to illustrate the effect of point of ignition on explosive performance in a blast hole. Its slow-explosive burning speed may be 1,000 to 3,000 feet (305 to 914 m) per second, so that it would take ⅕ to 
second for a 200-foot (61-m) borehole to fire. If ignited at the top, the upper rock might be moved a considerable distance before the bottom fired. In one way, this would act to lighten the bottom burden and help it to pull, but it might also serve to “uncork” the borehole and allow the bottom of the charge to blow upward rather than horizontally. On the other hand, the force of the upper part of the explosion, reacting against a heavy burden, might press down on the unexploded charge with great force and seal it.
FIGURE 9.21 Loading with instantaneous cap primer.
If the column were set off from the bottom by electricity, the toe would be well kicked out before the explosion reached the top. If several caps were placed at intervals in the column and fired together, the whole burden would be moved out at approximately the same time.
The same action is found in high explosives, although the rapid detonation makes it less effective. (See Fig. 9.21.) A 200-foot (61-m) column of a 40 percent dynamite with a velocity of 10,000 feet (3,050 m)per second, fired from the top, would explode at the bottom 
second afterward. If Primacord were used to the full depth, the detonation would take only 
second. If electric caps were used at top and bottom, the time at the bottom would be about 
second after the top, but the lag to the center would be second.
At first glance, these small fractions of a second might not seem significant. In many cases they are not, but sometimes they do have an important effect on both the performance and the concussion of a blast.
FIGURE 9.22 A 700-ton (636,000 kg) blast recorded by camera.
If long-borehole blasts do not give the desired effect in fragmentation, throw, or any other way, it may be advisable to change the location, number, or type of primers to try for better results.
Cameras can be obtained that will take pictures of the various stages of the explosion for study. These may be high-speed streak cameras, which can take 550,000 pictures per second, or instantaneous pictures, as seen in Fig. 9.22.
Precautions. Precautions to be observed in regard to a primer in placing and leaving it in the hole include: placing it in such a manner that it is not subject to shock or jar, and is not penetrated by rock splinters or other sharp objects; that it is not to be wet for a longer time than the powder, the cap, the fuse, or the wiring can stand; that the fuse and wires lead to the top without kinks, and are held so as not to be damaged by placing and tamping of additional charges and stemming.
Water Resistance. This is the resistance of an explosive to penetration by moisture, and/or its ability to explode when wet or damp. It is important when holes are wet, or when there is a long time between loading and firing.
Gelatin dynamite and slurry are good to excellent. Other dynamites are good with intact wrappings, fair to poor if cartridges are torn or punctured.
Ammonium nitrate has very poor resistance, but can be kept dry in plastic bags.
Electric caps are waterproof. Fuse and Primacord are practically waterproof except at cut ends, where fuse has no resistance. Cut Primacord absorbs water slowly. Damp sections are difficult to detonate.
Transporting. Large quantities of explosives should be transported in special vehicles marked in accordance with state or interstate laws. Smaller quantities may be carried in an ordinary car or truck, with any required warning signs made so that they can be removed when not in use.
Caps and explosives should be carried in different trips or vehicles unless quantities are small, in which case they may be carried in one vehicle if kept well separated and if permitted by law.
ICC regulations are accepted by most of the United States for intrastate transportation, but some have more restrictive laws.
Storage. Different classes of explosives should be kept in separate magazines. These should be far enough apart that an explosion in one would not affect the other; and should be surrounded so far as possible by earth barricades or higher ground so that the force of an explosion would be deflected upward.
Magazine areas should be as far as practicable from roads, railroads, or structures, and should be posted with warning signs and fenced if possible.
Magazines should be constructed of cohesive fire-resistant material, such as sheet iron, or soft material which will tear or crush rather than separate into flying fragments. Ventilation and protection from grass fires and from excessive heat should be provided. Doors should be heavy and provided with strong locks.
Portable magazines to hold a few cases of powder or boxes of caps are most easily made from large metal tool or packing boxes fastened with padlocks. When properly marked these are legal in most states, although many laws and regulations recommend more complicated units. When a portable magazine is to be left on a job, it should be chained and locked to a tree or other anchor.
Handling. Dynamite may cause severe headaches. This is especially apt to occur if it is unwrapped and handled with bare hands. Different brands and strengths differ in headache-producing qualities, and individual reaction is highly variable.
Persons handling explosives should not smoke and preferably should not carry matches. A complete list of safety precautions recommended by the manufacturers will be found in each box of dynamite. A complete list of the “don’ts” issued by the Institute of Makers of Explosives is included in the Appendix.
Boxes. Dynamite is usually packed in 50- or 60-pound (22.7- or 27.2-kg) boxes, although 25 pounds (11.4 kg) may be available. The wooden box, which was inseparable from blasting operations for many years, is no longer used.
The standard box is now made of fiberboard, with a full lift-up cover whose overlap provides double sides for the container. Sealing is by means of tape.
The box is lined by a polyethylene bag, which is an effective moisture and chemical barrier, and can be readily opened and reclosed in the field.
A small possibility of damage to both the dynamite and the liner, with resultant contamination of the fiberboard with explosive, is the reason that these boxes should not be burned, except with the same precautions as with dynamite itself.
Holes may be loaded in a number of ways. These may be classified as solid, string, spaced, deck, and spring.
Solid. In solid loading as much explosive is crammed into the hole as possible. Free-running explosive is poured into the hole, or blown in pneumatically from bulk carriers or from portable hoppers. The air tends to build up static electricity, so the unit must be grounded, and other precautions taken. (See Fig. 9.23.)
Water gels are usually loaded in their plastic containers, which may be intact or slit, but may be poured in. On large-scale operations it is economical to use a truck pumper, which may mix the ingredients also.
Cartridges smaller than the hole are slit up the sides in two to four places, so they will spread when tamped. Slitting is better than unwrapping, because of reduced danger from spilled powder or of headache from skin contact, and the wrapper ends prevent powder from sticking to the tamper.
Some cartridged explosives have special perforated wrappers that do not need slitting, as they tear and unwind under heavy end pressure.
Tamping. Tamping is a process of compacting explosives in boreholes by comparatively light blows, and/or pressure, of a stick or weight. This tool must not have exposed metal of any kind.
For best compaction, tamp each cartridge or layer separately with a firm, pressing stroke. Sharp blows are less effective, and should be avoided.
FIGURE 9.23 Grounding a loading hopper. Note: Mounted Du Pont Airloaders should be grounded to the ground with cable or wire. DO NOT use chain.
A tamping stick should be of round wood, with slightly smaller size than the smallest part of the hole, with a straight cut across the working end. If the hole is too deep for the use of a single tamping stick, several sticks should be drilled lengthwise and strung together with a cord. When the cord is slack, the stick will fold and can easily be fed into the hole. If any stick is held, and the cord tightened above it, the joints below the pull will be made rigid.
If the lower end of the stick wears to a taper, it should be cut back. The taper may punch holes in the tops of cartridges that would not be filled by pressure on the next one placed, it may grind some of the powder against the sides, and it may stick in cartridges and pick them up.
Large blast holes made by rotaries are usually tamped with a block on the end of a rope. The block should be of hardwood to resist abrasion, be slightly smaller than the bore of the hole, and have a flat end.
If weight is needed for heavy tamping, or working in wet holes that would float wood, the block may be drilled and weighted with lead or other heavy metal plugs, which should be covered with wood.
This type of block is not adapted to ramming down cartridges which have stuck in the hole above the bottom, as it may cause excessive side friction. A special block with a chisel-point stake that will break up the stuck cartridges is better.
These blocks are shown in Fig. 9.24.
Deep Holes. It is a good plan to check a deep large borehole before loading it, by inspecting it with a flashlight or sunlight reflected from a mirror, or sounding it with a tamping block, to make sure it is not obstructed. The block can be used to knock obstructing pieces or scale to the bottom.
Cartridges may be dropped into shallow holes or deep smooth ones. (See Fig. 9.25.) If the hole is deep and rough, and there is a possibility that they may stick partway down, they should be lowered. All explosives manufacturers will provide cartridges equipped with means to attach a lower rope. A band of special Scotch tape and a readily disengaged hook are used.
The impact of the cartridges on the bottom and the weight of the column above frequently compress charges well enough so that tamping is not necessary.
FIGURE 9.24 Tamping and cutting blocks.
FIGURE 9.25 Loading deep holes.
If the hole is ragged or partially caved so that it is not practical to load it with cartridges, a free-running explosive may be poured down it. If it blocks the hole and starts to build over an obstruction, it should be poked down with a long jointed pole or a dislodging block.
If such a hole is wet, it may be necessary to use a water-resistant slurry or water gel, either placed or poured.
A tamping block used for unwrapped dynamite should be kept clean by resting it on a box or some sacks when it is not in use.
String. If the borehole is wet enough that slit or unwrapped dynamite would be spoiled, or if solid loading would make too heavy a charge, cartridges somewhat smaller than the borehole, but not small enough to fit side by side, may be dropped in one after the other without tamping, or after having tamped the bottom cartridge or two.
This is the easiest way to load, and is satisfactory for small or occasional blasts. However, it is inefficient. More rock must be drilled than is necessary to hold the charge that is used. Part of the strength of the explosive is wasted in the air cushion around it.
Spaced. Spacers may be used to string cartridges out along a hole that is not to be fully loaded. These may be square, round, or hollow pieces of wood, tile, lean concrete, or rolled cardboard. They are usually made up ahead of time, in lengths of 8 to 10 inches (20.3 to 25.4 cm). There should be sufficient airspace around them to allow for detonating cord or wires, without squeezing or rubbing.
Spacers may be alternated with cartridges or pairs of cartridges in the parts of the borehole that are not to be fully loaded. The primer cartridge should have at least one additional stick in contact with it.
Decking. In large boreholes, charges which are to be strung out are usually separated by solid plugs of sand or other stemming material, and each section of the charge primed separately, unless fired with Primacord or other detonating fuse.
Stemming. Stemming is inert material such as dirt, sand, or finely crushed rock that is used to fill parts of a borehole that do not contain explosive.
Its primary use is in filling vertical holes from the top of the powder to the surface. Its use improves breakage by confining the force of the explosion, and adds to safety by preventing accidental igniting of the charge before it is fired.
Stemming is also used to space out charges, as in Fig. 9.25.
The minimum depth of top stemming should be about 1 foot (0.3 m) in a 1½-inch (3.8-cm) hole, and 12 feet (3.66 m) in a 12-inch (30.5-cm) hole. Deeper stemming is used where the top will be shattered simply by having its base blown out from under it. About two-thirds of a very high face may be broken in this way, with explosive in only the bottom third.
Carrying the load too high in a hole is at best a waste of powder, and at worst can mean excessive and damaging rock throw and noise. On the other hand, too little explosive and too much stemming will give poor top breakage.
Drill cuttings are often the best source of stemming, particularly from rotary drills or air drills equipped with dust collectors. It is important that stemming contain no sharp pieces or sizable stones that might cut wires or fly in bullet fashion. Moist material usually is more effective than loose dry sand.
Mines and quarries that have rock-crushing plants and that use large-diameter holes often use fine crushed rock (screenings) for stemming. This may be hauled in Dumpcrete trucks that can dump directly in the hole through chutes. Ordinary dump trucks may leave a pile of screenings near each hole or group of holes, for distribution by hand shovel or wheelbarrow.
The length of fuse determines the time which will elapse between lighting it and the explosion. Regular sequences of firing can be obtained by varying the lengths of fuse in different holes, and lighting them at the same time.
Fuse does not light readily with a match because of the small area of powder exposed, and the likelihood of wax from the coverings being spread over the end while it is being cut.
If only a few fuses are to be lit, good results can be obtained by splitting each fuse with a knife or razor blade, as in Fig. 9.26, bending the fuse so the opening is down. It may then be lit with a match. Care should be taken to keep the fingers out of a line with the end of the fuse, as it will spit out a jet of flame.
The split fuse can be ignited more readily by having the opening horizontal or upward, placing a broken-off match head between the halves, and squeezing them together, as in (B). The match head gives a much hotter flame than the stick.
It is possible to buy a number of devices which simplify the lighting of fuses. The match lighter is a short paper tube which fits over the end of the unslit fuse and is coated on one end with a compound similar to that of a safety match head. This is readily ignited with a match or the edge of a match box, and subjects the fuse end to intense heat.
FIGURE 9.26 Lighting a fuse with a match.
This lighter throws a jet of flame which resembles that caused by ignition of the fuse. To tell whether the fuse is actually burning, it is necessary to observe a moment later whether a thin stream of smoke is issuing from it. If it is not smoking, the lighter should be removed and another applied.
The pull-wire lighter is of similar material, but clamps on the fuse and is ignited by pulling a wire.
The lead spitter is a coil of thin lead tubing containing black powder. A piece is cut off, lit with a match, and the resulting hot flame used to then light the fuses. The lead melts back as the powder burns.
The hot wire is similar to a fireworks sparkler. It burns slowly with a very hot flame, and is the safest and most dependable of the devices listed.
A burning fuse will light black powder by contact. When used to explode dynamite, it must be connected to an explosive cap.
Electrical firing requires a complete circuit from the power source through all the caps and back to the power source. (See Fig. 9.27.) One cap, or several hundred, may be used.
A cap may detonate the charge through a primer in the borehole, or initiate a detonating cord hooked up with a number of holes.
Source of Current. One standard source of electric energy for blasting is a blasting machine. This may be a generator which delivers high voltage when a handle is pushed or twisted vigorously.
These devices are rated according to the number of caps in series which they can fire at one time. Rating is usually conservative. Their efficiency should be tested occasionally, particularly if they are not in steady use, as they may deteriorate rapidly in damp storage. Testing is done by means of a special rheostat which sets up a resistance in the line; and from one to four blasting caps. If the machine will overcome the rheostat resistance of its rated capacity, and fire the caps in addition, it is in good condition.
Newer-type blasting machines operate on the capacitor or condenser principle. Current is supplied by flashlight-type dry cells, or by a 12-volt rechargeable nickel-cadmium unit. This modest current is built up to high voltage in condensers, then discharged into the blasting circuit.
The high voltage may be built up when required, by pressing a Charge button or switch, then used by depressing a Fire button, after a buildup time of 5 to 30 seconds.
Some models automatically build up voltage between uses, and can fire instantly.
FIGURE 9.27 Detonating cord layouts.
Storage or dry cell batteries may be used in emergencies for shooting a small number of caps, but they are not considered to be adequate or safe for regular use.
High Lines. Where electric lines (called high lines) are available, 220 to 440 volts may be used for firing. Special switches are made for connecting into such lines. They automatically shunt or short out the firing lines until the moment the switch is pulled.
Series. There are three basic types of circuit—series, parallel, and parallel series.
When the caps are arranged in series, Fig. 9.28, the current must have enough force, or voltage, to overcome in succession the resistance of the lead wire, the caps and their wires, and the return lead wire, in addition to the variable resistance offered by connections between wires.
The voltage required can be calculated by Ohm’s law. This basic law states that the current, in amperes, in an electric circuit will be equal to the potential, or voltage, of the power supply divided by the resistance, in ohms, of the circuit. That is, if sufficient current is supplied at 110 volts to a circuit with 10-ohm resistance, the flow of current will be 11 amperes. If the voltage is 6, the flow will be only .6 ampere.
A single cap requires a current of about .5 ampere. A series of caps takes 1.5 amperes, with sufficient voltage to overcome all resistances in the circuit.
The tables in Fig. 9.29 indicate the resistance of caps and wires commonly used. The supply of current should be well over the calculated need, however. Minute differences in the bridge wires in caps may vary their resistance, so that a weak current might burn some of them through and break the circuit before all are exploded. Be specially liberal if the series includes both regular and delay caps.
Series circuits are easy to lay out, to hook up, and to test.
Caps made by different manufacturers must not be used in a series, because of variation in current requirement for detonation.
FIGURE 9.29 Resistance of caps and wires. (From the 16th Edition of Blasters’ Handbook.)
Parallel. In a parallel circuit, Fig. 9.30, the current does not go through the caps one after another but goes through all of them at the same time. A poor connection on a cap wire affects only that cap. The voltage requirement is lower than when the same number of caps is shot in series, but more amperage is needed.
Two caps in series have twice the resistance of one cap. Two in parallel have only one-half the resistance of one, as less potential is required to force the current through a large conductor than a small one. But where 1.5 amperes was sufficient current to shoot a whole series of caps, .5 ampere is required for each cap placed in parallel.
Parallel wiring is therefore preferred where a source with low voltage and high amperage, such as a storage battery, is to be used. It is not suitable for blasting machines, and the results with dry cell batteries are doubtful. High lines are equally efficient with either arrangement.
The most common simple parallel hookup is the second one shown, and it is not recognized as such by many who use it. It is the most convenient way to fire a small irregular group of blasts.
In figuring a parallel circuit, the resistance on one cap is divided by the total number of caps, and in a large blast, may be so small a figure that it can be ignored. The resistance of the two bus wires, between the leads and the last cap, is approximately one-half the resistance of the same length of the same wire used for a lead. Lead wire resistance is the same as in a series circuit.
The lowered resistance of the bus wires is due to the fact that some of the current is diverted at each cap. Full current is present at the beginning, and zero current at the end, so that it averages out to about one-half current for the full length of these wires.
Parallel Series. This layout, Fig. 9.31, makes it possible to shoot large numbers of caps without requiring excessive voltage or amperage.
There is some disagreement about the balancing of the size of the different series. Technically, each series should have the same number of caps. Many blasters, however, claim better results when the series differ from each other by a set amount. This is said to be particularly advantageous when firing an excessive number of caps with a blasting machine.
When the series are equal, and juice is put in the line, all caps are equally heated and should detonate simultaneously. If any series gets current but less than its share, due to a poor connection or other defects, it may not fire. However, as the other series fire, current will cease to go through them, and all of it will go through the remaining one, unless the wires are broken, until it fires also.
If the series have different numbers of caps, at first the current will flow most strongly through the series with the fewest caps and least resistance, and having fired that, will concentrate on the next longer string, and so on through the whole group. If the current is weak, there may be a brief but definite time interval between the series, so that short strings should be at the face. If it is strong, explosions may be simultaneous.
As the current from a blasting machine flows very briefly, it is possible that in a graded pattern only the shorter series would fire.
In most blasting patterns, the equal series hookup will be very much simpler than the graded series, which is generally considered obsolete.
Making Connections. It is most convenient to have the cap wires of such length that they meet each other with a moderate amount of slack between the holes. If they are short, an extra piece of wire must be spliced in at each connection; if they just reach, insulation or primers may be strained while splicing; and if much too long, they will need cutting, or will make a tangle of loose wire that may lead to mistakes in connecting, and accidents.
While connections are being made, the power or far end of the lead wires should be fastened together to ground out any induced current. The electrical source should be locked or at least removed from the immediate vicinity of the wires.
In a series circuit, the current runs from one lead wire through all the caps and their wires, to the other lead wire. The insulation on each lead is stripped back with a knife or plucked off with pliers for an inch or two, and the wire is bent to a tight loop.
The cap wires are pulled apart at their soldered connection, and each wire is connected to one from an adjoining cap. When all the caps are connected in this manner, the two end wires are connected to the lead. Figure 9.32 (A) to (C) shows connections commonly used between caps, and Fig. 9.32(D) shows two cap-wire-to-lead-wire hookups.
It is usually possible to arrange the circuit so that it is convenient to hook both end caps to the lead. However, in a one-row straight-line layout, an extra wire must be used to connect the last cap to the lead.
If cap wires are long, such connecting wires are usually made up of scrap wire left from preceding blasts. It is good practice to extend the leading wires some distance with a lighter and less expensive wire, to reduce damage to the ends. If not enough surplus wire is available, connecting wire may be brought on the job on a spool and cut as needed.
The use of scrap wire involves a risk of misfires due to breaks inside the insulation, which is not justified by the small value of the wire saved.
When the last cap has been connected, the whole series should be rechecked, to make sure that no hole has been omitted, that no loose ends of wire are lying around, and that connections are tight. It is good practice to squeeze each connection with pliers at this time.
Bare connections may be propped up on sticks or rocks where necessary to keep them out of water, or from contacting wet ground. If any connections are unavoidably wet, they may be tightly taped or smeared with water-resistant grease.
If the shot is only of a few caps in a limited area, and the electrical source is of ample power, precautions against bare wet joints are necessary only if the water has a high mineral content.
Electrical Hazards. Electric caps are supplied with a connection between the ends of the wires which is called a shunt. This prevents the accidental buildup of opposite electric charges in the two wires, which might pass enough current through the cap to explode it.
Such charges may be caused by the near presence of electrical machinery or transmission lines, a radio transmitter, stray currents in the ground, thunderstorms, or static electricity from dust storms or escaping steam. Such currents in lead wires may often be detected by inserting a No. 47 radio pilot lamp in the circuit instead of the cap. If it glows, conditions are unsafe.
The best precaution to take in blasting near an electrical hazard is to use fuse caps and Primacord. However, certain precautions can be taken which will reduce the danger of using electric caps.
Lead or other wires should not run parallel to electric lines.
The shunts should be left on cap wires until they are connected into the blasting circuit, and the circuit should be shorted until ready to fire.
The two-cap layout in Fig. 9.33 illustrates the method to be used. The lead wire is shorted at each end. A wire is connected to one lead, and to one leg of a cap. A connection is made from the other leg or the shunt, to one leg of the next cap which has its other leg connected to the return lead. Making cuts where indicated will include the caps in a firing circuit which is closed until the leads are separated at the battery end.
FIGURE 9.33 Protecting caps against stray currents.
Even with these precautions, however, blasting should be discontinued if there is a thunderstorm within 5 miles (8.0 km), or other severe hazards exist.
Crackling in a portable battery-type radio (not FM), left at high volume, provides warning of approach of a thunderstorm.
Testing. A circuit tester should be used for checking before attempting to blast. This device consists of a galvanometer and a silver chloride dry cell, which produces a current too weak to fire a standard blasting cap. The lead or cap wires are fastened to its terminals, and the action of the indicator needle shows the condition of the circuit.
If the circuit is good, the indicator needle will move an amount inversely proportional to the resistance offered by the caps and wire used. If the needle does not move, there is an open break. If it moves only slightly, a loose connection, or a break with wires just touching, is indicated. If the needle moves farther than it should, a short or a ground is present.
Each hole may be tested before hooking into the circuit, and a single test may be made from the power end of the lead wires when wiring is complete.
Any trouble in the system can be spotted by making a series of tests with a long connecting wire. In Fig. 9.34, the connecting wire N is fastened to one lead and to a tester post. The other post, or a wire securely connected to it, is touched in succession to connections E, F, G, H, and C. The bad reading will show up whenever the difficulty is inside the circuit being tested. As an example, if normal readings are shown at E and F, but an abnormal one is found at G, the trouble is in number 3 cap or its wiring.
FIGURE 9.34 Testing for break in blasting circuit.
After the blasting circuit tests properly, an additional test is made at the power end of the leading wires.
Warning. Warning must be given of intention to blast. The type of warning may be determined by either law or custom. For large blasts, particularly in pits employing numerous workers and machines, blasting should be done at specified times, as at 12 and at quitting time, and should be preceded by definite and well-understood signals, such as horns, sirens, whistles, or yelling, long enough in advance to notify all persons and give them time to prepare.
A usual blast signal is a call of “Fire” or “Fire in the hole,” repeated two or three times at intervals of 10 to 30 seconds. Signals should be arranged by which workers detailed to watch entrances to the area can stop the blast if necessary.
It is the responsibility of the blasting crew to make sure that all personnel are out of the way, machines are protected, and no visitors or trespassers are where they can get hurt.
The area, and particularly any roads or paths leading into it, should be marked with warning signs. If any public roads are within the danger area, traffic should be stopped at a safe distance.
Firing. During the signaling, the wires are connected to the blasting machine or switch, or if a battery is used, one wire is connected to a post. To fire, the blasting machine handle is slammed to the bottom, or the switch is closed.
There is a wide divergence of opinion as to what constitutes a safe distance from which to fire a blast. Some experienced people will take shelter behind a nearby rock or tree, whereas others consider 500 feet (152 m) a bare minimum. No one should stand in front of a rock face, at any distance.
Proper barricading may be as safe as distance and more convenient. Full protection requires some sort of roof or overhang. A very safe spot is in the tucked-in bucket of a big front shovel which is turned away from the blast. The shovel itself may be protected by wood lagging on the rear of the cab.
The return to the blast should be slow for several reasons. The fumes, which dissipate in a few minutes in the open, are toxic and may cause severe headaches and nausea. If more than one hole has been shot, one of them may fire late. Rocks are occasionally thrown so high that they take a long time returning. Rocks or debris may be lodged precariously in trees.
Light blasting, Fig. 9.35, includes loosening up of shallow or small outcrops of rock and breaking boulders. It may constitute the entire job, be done in connection with dirt excavation, or follow heavy blasting which has failed to cut to grade or slope lines or has left chunks too large to load.
Chip Blasting. Shallow rock outcrops are most conveniently broken up by drilling and blasting. Unless the rock breaks readily along planes more or less parallel with the surface desired, it will be necessary either to drill much deeper than grade or to space the holes closely. It is often good practice to blast each row before drilling the next.
Loading may be light, or very heavy, but in general it is necessary to use more powder per yard of solid rock than in heavier work.
Laminated or jointed formations may be shaken apart by light charges. Fragments may be thrown long distances, and mats used to confine them are more subject to damage than with deeper blasts.
The amount and direction of throw can often be controlled to a large extent by drilling and loading procedures. A vertical hole causes scattering in all directions. A sloped hole tends to leave the lower slope in place and to throw the upper one away from it. Throw is reduced by increasing the number of holes, reducing the charges, or drilling deeper than required by the breakage line. These last two place the powder deeper, where more of its power is applied to breaking and less to scattering.
When breaking must be done exactly to a line, holes are drilled closely along the line and a variable number left without charges, as discussed on page 9.12.
Blockholing. Boulders and oversize pieces of blasted rock may be broken by drilling a hole slightly more than halfway through, and exploding a small charge of dynamite in the hole.
Fragments may be thrown for long distances, so that protection should be provided for the blaster and other personnel. High-velocity explosive, or large charges, will produce finest fragmentation.
Chip blasting may be called blockholing also.
Mudcapping. Ledges may be chipped and boulders broken by mudcapping instead of drilling. Heavy charges of dynamite, preferably of the highest-velocity type that is obtainable, are laid on the surface of the rock, primed, and covered with a few inches of mud. The explosion acts as a giant hammer blow and should split or crush the stone.
Knowledge of the grain and jointing of rock is important in successful mudcapping. The charge is placed in the same place which, in hand breaking, would be hit with a hammer or opened with a wedge. In general, hammerlike crushing is most effective on loose boulders, and splitting on ledges.
It is a common error to suppose that the force of black powder is chiefly exerted upward, and that of dynamite downward. In each case the explosion acts equally in all directions, but when it acts slowly, it can find and follow paths of least resistance, where the quicker-acting dynamites deliver such a rapid blow that they will crush objects under them, even when not confined. However, a study of the table in Fig. 9.36, showing quantities of dynamite used for blockholing and mudcapping, will show the waste involved in open explosions.
FIGURE 9.36 Charges for boulder blasting.
The mud pack over the charge is usually 2 to 6 inches (5.1 to 15.2 cm) thick. It serves to confine the explosion slightly, increasing the force exerted on the rock and reducing noise and airborne concussion. Mud is preferred to any other substance. It is much more effective at confining the explosion than dry or damp dirt or sand, as it packs and sticks together better. It should be free of stone or pebbles that would create a hazard by flying long distances.
Charges can be fired on bare rock, but they are less efficient and even noisier.
Mudcapping is wasteful of powder, excessively noisy, and less certain in effect than drill hole blasting. However, it causes less rock scatter than other methods of shallow blasting, and does not require the presence of a compressor.
Snakeholing. Boulders are most readily broken if they are lying on the surface of the ground. If partly buried, the earth or other rock around them provides a support and a vibration-absorbing cushion that may prevent or reduce breakage.
Embedded boulders may resist machinery that can handle them readily once they have been loosened up.
Snakeholing consists of making a hole beside or under a boulder, and firing a charge sufficient to roll it out of the ground, and preferably to break it also. Any further breakage required can then be accomplished by mudcapping.
Snakeholing is more laborious than mudcapping, but is more economical of powder and is much less noisy.
Rock is broken or fractured by an explosion in three ways: compression, shear, and tension. Compression is obtained by the direct, hammerlike blow of the explosion against an unyielding rock mass. An explosive that is so deeply buried that it cannot break out to the surface breaks by compression only. This is the least effective way to use it.
Shear is movement of pieces or blocks of rock along lines of weakness. Tension is produced by reflection of the explosion back from an unconfined surface or face of the rock.
In a hard rock, maximum effectiveness of explosives is in tension. Tensile strength is only about one-tenth of shear strength, and shear strength is only one-tenth of compressive strength. This means that a blast that can break out to a free face at an efficient distance may produce 100 times the rock breakage of one that is completely confined.
The importance of tension has been demonstrated by setting off a tightly confined explosive far enough below a horizontal surface that the explosion could not break out, but near enough that surface rock was shattered in a cone-shaped crater, separated from the small explosion chamber by solid rock.
It is very fortunate that rock breakage is not produced mostly by the outward movement of gases from an explosion, as in this case rock throw would be many times greater than it is. In a well-engineered and successful blast, most of the rock moves less by explosive force while being broken than it does afterward by slumping under the pull of gravity.
The improved fragmentation obtained by using millisecond-delay caps is partly due to creation of a series of free faces from which waves of succeeding explosions reflect to produce tension in the rock.
One of the contractor’s problems in connection with blasting is the possibility of real or imaginary damage being done to structures in the vicinity.
An explosive, if properly used, will expend most of its energy in shattering the rock immediately around it. The remaining energy will set up waves or vibrations in the ground, and sound and concussion in the air.
Noise. The noise of an explosion may cause most or all of the neighborhood difficulties. Mudcaps, shallow blasts, overloaded holes, fractured rock, and other conditions that allow the explosion to break out into open air before expending its energies, are the cause of complaints all out of proportion to the amount of explosive used.
In the first place, the noise attracts attention to the fact that blasting is going on. It causes the householder to concentrate on trying to feel the jar or shake of the blast, to look for cracks in plaster, and to speculate about other damages that might be done. In many cases, the sound of the blasting will annoy sensitive people so that they will invent or exaggerate physical effects. The contractor or quarry operator’s first rule is therefore to blast as quietly as possible in any area where there is a possibility of complaint.
This means a first rule of no mudcapping. This technique is not only wasteful of explosive, but a sure way to lose the goodwill of the neighborhood and of the insurance adjuster.
Boulders and oversize blast fragments should be drilled before blasting. The noise is tremendously reduced, and it will usually be found that the saving on explosive and the better fragmentation obtained will more than outweigh the cost of the drilling. When there are only a few pieces, the nuisance of clearing the pit for blasting may be avoided by plug-and-feather splitting. In brittle rock, a crane with a skull-cracker steel ball may be the most economical solution.
To minimize noise, do blasting when weather conditions are more favorable, that is, clear to partly cloudy skies and relatively warm daytime temperatures. Unfavorable weather conditions are those when the air is relatively still. These occur when the days are foggy, hazy, or smoky and the temperature is constant from the ground up to a high altitude. About the only general rule is that the blaster should consider avoidance of noise one of the important objectives.
Sound travels rather slowly. Its distribution is affected by winds, as shown in Fig. 9.37; by reflection from hills, clouds, or atmospheric layers; and by temperature and humidity.
Characteristics of Damage. In most cases, there is an observable difference between building damage caused solely by vibration, and that resulting only from structural defects and/or settlement that change the shape of the building. Most of the distinction is in the pattern of wall cracks.
Cracks caused by vibration are usually fairly uniform in width with little or no displacement of the sides relative to each other. Vibration cracks tend to fan out diagonally from corners of windows and door frames.
FIGURE 9.37 Sound travel from blast.
Either type of stress frequently causes walls to crack along their junctions with a chimney or column. Settlement or tilting of the chimney only, or the building only, usually results in an opened or spread crack on one side, and a closed or even pressure-squeezed crack on the other. Vibration more often produces equal-width cracks on both sides.
A distinction can sometimes be made on an age basis. New cracks show clean surfaces of material; older ones become progeressively grimier, the rate depending on conditions. Dust, spot deposits, and light-bleaching of color pigments are indicators. Recent cracks should be clean and unfaded.
Concussion. Airborne concussion is responsible for a large share of the damage in bombing, and in accidental detonation of explosives, but is rarely a factor of importance in blast damage. It consists of one or more waves of highly compressed air moving outward from the explosion.
Sufficient explosive to cause concussion more than a few feet (meters) away should not be used in mudcapping. Even very heavy blasts in solid rock cause little or no concussion if they are laid out and loaded properly.
Any damage caused by concussion is usually obvious. Glass breakage in closed windows at right angles to the path of the waves is the most common result. In the absence of extensive glass breakage, it is very doubtful that any other parts of a structure could be damaged.
If a building is to be endangered by blasting, windows should be opened or removed, particularly those on the facing and far sides of the house. Store windows may be braced as a routine precaution. Careful check should be made of the condition of plaster and masonry, so that claims need not be paid on preexisting defects.
Rock Throw. Unexpected damage may be done by rock or other material thrown through the air by blasts. In general, shallow blasts, overloaded holes, shots in rock with irregular resistance, and block-holed boulders give the most trouble in proportion to the amount of powder used.
Thrown objects may cause injury or death, and their control is therefore of first importance. Property damage may or may not be severe, but at least claims filed on this ground are usually sincere.
Danger of damage from rock throw may be reduced by increasing the number of holes so that smaller charges may be used, by sloping holes to throw rock away from danger points; by reducing the quantity or strength of powder, and handling any resulting oversize fragments by blockholing under mats or by the use of larger machinery.
Covered Blasts. Throw can be closely controlled by working downward, using small blasts and covering them with mats or chained logs. If the cover is large and heavy in proportion to the strength of the explosion, it will prevent any scattering of fragments. If the charge is heavy enough to lift the cover, it will move somewhat less than the average distance of throw to be expected from an uncovered blast, and fragments with higher-than-usual velocity will be held in.
It is important that the cover extend several feet beyond the area being shot, particularly if the charge is heavy enough to lift the mat, as fragments might escape under its edges.
When a power shovel is used to remove the shot rock, it is advantageous to use a woven steel mat as it is easily handled with chains, and provides a quicker and more secure cover than logs. The mat is lowered over the holes, or dragged in such a manner that it will not damage the wiring and cause misfires.
Logs are used when no mat is available, or when there is no machinery on the job which can handle one. They should be long enough to overlap the blast at both sides, and light enough to allow the crew to carry them by hand. Two chains should be laid on the ground first, the logs piled, and the chains fastened over them, preferably by wired square knots.
Chaining is important, as unfastened logs may be thrown farther than rocks.
Neither mats nor logs should be laid directly over mudcaps as they are liable to be thrown long distances and be severely damaged as well.
Blasting mats should be used wherever there is the slightly possibility of fragments reaching people or property. Even a scattering of sand or fine pebbles on their property will make people nervous and resentful, and is an indication that loading should be reduced or the technique changed to a safer one.
Reasons for Complaints. Studies indicate that much greater vibrations are produced in house structures by slamming doors, running, and often street traffic, than by even severe blasting. It would appear that at ordinary distances the ground and air vibrations set up by heavy blasts are so weak as to be incapable of affecting any structure. Yet complaints and claims for damage pour in on every blasting job. Why?
There are a number of reasons. One is that the ability of rock, soil, and water to transmit vibration varies much more than is indicated by the relatively superficial testing that has been done.
It would appear from an old table of recordings that 600 pounds (272 kg) of explosive would not produce sufficient ground waves to damage a house on average overburden 100 feet (30.5 m) away (!!). But there are records of high five-figure awards paid for damages done to a village 2 miles (3.2 km) away from an underwater blast of this size, indicating a difference of over 10,000 percent between theory and fact.
Contractors frequently blast much more heavily than is indicated by their records and statements, particularly when a job gets behind schedule. Mistakes in loading can occur. Variation in the strength and quality of explosives can be a factor.
Most of the checking of blast damage to date has been done by representatives of mining and insurance interests, who are more interested in disproving it than in making an impartial study. Some of the instruments used for measurement leave much to be desired.
There are also psychological reasons for exaggeration of blast damage. Bomb damage received much publicity in modern wars and recent terrorist attacks and has made people overly conscious of the dangers of explosives. There is also fear and resentment of the unusual, which makes blast vibration appear more significant than that from a truck.
Prestressing. One of the principal defenses advanced by defendants in blasting damage suits is criticism of the condition of the structure before the blast. If it is in a condition of stress due to unequal settlement, warping or shrinking of timbers, or overloading, it will change in shape and its plaster will crack.
If a blast vibration is within “safe” limits, an overstressed condition may cause cracking from the blast. The theory is that if the blast had not been set off, the same cracks might have developed shortly from natural causes.
In general, the poorer the quality of construction, the greater the probability that stresses will develop, plaster crack, and misalignment occur. But this is not always so.
The prestress argument is unpleasantly reminiscent of the whitewash given the Donora smog by a group of doctors. They said in effect that it was nothing to fuss about, as only people with a previous history of respiratory disease had died.
It would be unjust to allow reckless blasters to evade payment of damages on these grounds, or to make property owners go without recompense because their building standards fall short of those set up by the U.S. Bureau of Mines. However, blasters should not be compelled to subsidize substandard construction. It is likely that most cases where prestressing is actually proved should be subject to compromise settlements.
Water Supply. Blasting sometimes causes springs and even deep wells to go dry. The vibration causes underground movements that may close water passages or open new ones. However, explosives probably are responsible for only a fraction of the difficulties for which they are blamed.
Underground water circulation is under constant change. Old seepage veins become plugged with mineral deposits, new ones are opened by solution and erosion. Changes in rainfall pattern, in conversion of forest land to farms, or back again, may alter the quantity and location of underground water over a wide area. Overpumping will lower the water table.
A new well may tap into an underground reservoir of limited size, which once pumped out will not refill. Such a well may show a very high yield on its first test but decline markedly after long use, when it comes to depend on circulating water only.
Keeping Out of Trouble. Under all ordinary circumstances, blasting should be kept light enough not to damage buildings. The job should be figured on a basis of conservative blasting, and the work done the same way.
Short-period delays are a real friend to the person who wants heavy blasts, but is surrounded by structures. Up to 70 percent of the charge of an instantaneous blast can be used with each of 10 or 12 short delay periods, without increasing the vibration. Or to look at it another way, the same loading can be used as for an instantaneous shot, and 10 periods used to cut the damage potential by four-fifths.
Even with ultraconservative procedures, inspections should be made of nearby buildings before blasting. If property is valuable, vibration-testing devices will be supplied by the insurance company or by a blasting consultant to measure the disturbance caused. Such instruments should be used to check the next blast in any building or area from which complaints are received.
As detailed before, noise should be kept to a minimum. If there are few people in the area, it should be possible to notify them before blasts, so that they will not feel it necessary to be tense all day waiting for an explosion. Another method, applicable to heavily populated areas also, is to set a definite time or times each day for shooting, and stick to it.
If a claim is made and is justified, it should of course be paid. But if it is clearly unjustified, it probably should not be paid even if apparently too small to be worth arguing about. One paid claim is likely to bring in a dozen or a hundred others, and the contractor might be forced to replaster and decorate a whole town before he or she knew it. Payment of any claim makes any other much harder to defend in front of a jury.
Of course, contractors should protect themselves with insurance, and usually do so. But in the long run the premiums they pay are based on what the company pays out for damages, so their interests are identical.
Tunnels are underground passageways of any size, and may be natural (as in limestone caverns) or made by animals or humans. Those discussed in this section are made by people. They serve a variety of purposes, including mining, water supply and drainage, laying sewer and other pipes, railroad and vehicular shortcuts or water crossings, and air raid protection.
Rock tunnels are driven through solid material that usually requires blasting and may support itself permanently, or at least long enough to allow setting up of bracing after digging out a short section. Soft ground tunnels involve digging or pushing aside soil, and the roof (called the crown) and the walls may require support before removing the soil. Mixed-face tunnels go through both types of ground, either together or in different sections.
People have driven tunnels since prehistoric times. They usually worked in rock, because the difficulty of digging it was more than compensated for by its ability to hold itself up. Cutting was done with hand tools, or by heating the face with wood fires, then throwing cold water or cold water and vinegar on it to cause sections to crack off.
The vinegar technique, with little or no ventilation, must have been really rough on the workers. A rough approximation of the atmosphere might be obtained by building a good blaze in a fireplace, shutting off the chimney damper, then putting out the fire with vinegar.
Layout and Problems. The methods used to drive a tunnel vary tremendously with the nature and water content of the material to be penetrated, depth and size required, surface conditions along the route, time allowed, and background of the workers doing the job. There is space in this section to indicate only a few of the problems most often encountered.
The diagrams in Figs. 9.38 and 9.39 show the layout of a simple tunnel job, and the names for some of its parts. If it is driven more or less horizontally into a hillside, the opening is called the portal. The working face, where the digging is done, is the heading. Vertical access tunnels descending from the surface to the main tunnel level are known as shafts.
In the tunnel itself, the floor is the invert, and the roof is the crown. The spring line is the meeting of the vertical sidewall with the curve of the roof arch. A supporting shelf cut at this line is the hitch. A small pioneer or accessory tunnel is called a drift. Standard cross sections are rectangular, round, and horsehoe-shaped.
FIGURE 9.39 Tunnel cross sections.
There are a great many special problems connected with even a simple tunnel project. To the open-cut worker, one of the most impressive is lack of space. Many tunnels have been driven with cross sections as small as 4 × 4 feet (1.2 × 1.2 m)—not even big enough to stand in.
Under some conditions tunneling can be done by handwork. It is not suitable where the tunnel is too small for man entry or the tunnel is so large that there is too much material to excavate. The size range for possible hand mining is generally accepted as 42 to 60 inch (1.1 to 1.5 m) height.
In the United States case of Washington, DC, which involved extending sewer lines over the top of the Metro Transit tunnels, where the shoring materials and other unknown materials were in the sewer alignment, hand mining is quite appropriate.
The possibility of hand tunneling is normally limited to reasonable ground conditions in the following soil types: rock, clay, and stiff sand and gravels. Some ground water can be tolerated, especially if the tunnel is headed uphill.
In appropriate ground conditions and with practically all pipe products, hand tunneling is safer and more efficient when it is done within a shield. The shield not only protects the workers but also provides for more directional control of the heading. Many shields can be steered mechanically or with hydraulic power.
For a relatively small pipe installation, such as a 36-inch (0.91-m) sewer pipe, and adverse conditions, such as high groundwater, the microtunneling technique may be used. This technique eliminates the need for open trench excavation and possibly tight sheathing with the need for a dewatering system. It is sometimes called trenchless technology.
The technique makes use of a microtunneling machine that uses a pipe jacking apparatus that guides the pipe through the ground as it advances. The apparatus is remotely operated with laser guidance to keep it in line. The microtunneling machine has a sealed head in the cutting face to keep the soil pressure balanced between the slurry injection system and the removal rate of the soil dug by the cutting head. The slurry injected through hoses from the back end of the machine and through holes in the front end of the pipe is generally a bentonitic mixture of earth material that lubricates the movement of the pipe and seals the outer sides of the pipe when it becomes stationary in its final position.
There are several problems that occur for the microtunneling machine. Because of its sealed head arrangement it can not deal with obstacles in the way of the advancing pipe the way a horizontal boring machine can. But the boring machine needs to be operated with manpower inside and other complications. Also there is a limit to the toughness of the ground that the pipe is driven through using microtunneling. Generally, if the ground is no harder than 5,000 psi (350 kg/sq.cm) resistance, then the microtunneling machine can operate.
A microtunneling machine performs better as the moisture content of the soil increases and the soil type changes from a stiff clay to a more granular material. In good conditions this method may be able to advance from 20 to 80 feet (6.1 to 24.4 m) per day.
Twenty- to 30-foot (6.1- to 9.1-m) diameter tunnels are big, yet they provide a floor width that would be considered skimpy for a haul road on top.
Equipment to be managed at and near a tunnel heading may include a drill jumbo (a movable frame almost as big as the tunnel, carrying a battery of drills), a machine for loading muck (the below-ground name for spoil) and railcars or rubber-tire trucks to remove it; the same or other cars or trucks to bring drill steel, bits, explosives, and other supplies to the heading; a locomotive to push and pull cars; and a switching or passing device to permit hauling units to get past each other, although often there is room only for a single width of track or roadway.
There will be high-pressure air pipes to supply the drills, and often large low-pressure ducts for ventilation. Overhead wires or ground cables carry electricity for light, power, and blasting juice. Water under pressure may be supplied for wet drilling. A system of drainage, pumping, or both may have to handle tremendous volumes of water.
In addition to the regular equipment there may be need for a diamond drill to make test and grouting holes, grouting equipment to seal off leaks and solidify wet ground, and/or a movable buffer to confine rock throw from blasts.
If the tunnel is to be timbered for support, or lined for support or for permanent use, the crews and materials for this work may follow the digging closely, and in any event will have to work in and over the single entranceway.
If driving is from a shaft, its bottom is another crowded point. Haulage equipment may be lifted to the top to dump, or it may dump into containers at the bottom. Supplies must be unloaded from the elevator cages and reloaded for hauling to the face. Crews leaving work, and their supervisors and inspectors, wait here for transportation to the surface. Pumps, compressors, and even drill and repair shops may be located in skimpy quarters excavated near the shaft.
Sequences are very exacting. The tunnel cycle (the succession of drilling, shooting, and mucking) must keep the largest possible number of workers and machines usefully employed, and the time interval from any operation to its repetition should not vary. Whenever possible, two or more operations should be performed simultaneously, as drilling the top of a face while digging the bottom, and installing lining a few feet (meters) back at the same time.
When two headings driven from one shaft are close together, one may be drilled while the other is shot and mucked. With increasing distance, the advantages of this arrangement are reduced.
Most tunnel crews are of the universal type, and perform all operations in the cycle. This saves the contractor from paying a crew waiting time because of a delay in a prior operation.
Speed. Under favorable conditions, tunneling may progress very rapidly. The Owens River Gorge Power Tunnels in the Los Angeles water system were driven as fast as 104 feet (31.7 m) in a day, and 2,442 feet (744 m) in the best 31-day period. Other tunnels excavated by using tunnel-boring machines, called TBMs, working in fairly ideal ground conditions have made even better progress.
Gold mines at Kimberley, South Africa, hold the depth record at 9,000 feet (2,743 m). These tunnels must be air-conditioned, as otherwise the heat would make it impossible to work in them.
The 12-mile (19.3-km) Simplon tunnel in the Alps is 7,000 feet (2,134 m) beneath the surface at one point. Temperatures up to 131°F (55°C) were encountered in drilling it.
A record for its time was established in twin power tunnels at Niagara Falls in Canada. These are each 51 feet (15.5 m) in diameter and 5½ miles (8.85 km) long. Together they required over 5,000,000 yards (3,830,000 cu.m) of excavation.
There is such constant improvement in tunnel-driving techniques, and increase in confidence to undertake bigger projects, that some or all of these records may have been surpassed by the time this book is in print.
Plant. The plant at a tunnel may include the tower, hoist, and hopper; compressors, low-pressure ventilation system, water pumps, electric transformers or generators, change rooms with showers and lockers, provision for emergency treatment of injuries, a blacksmith, forging, and bit dressing shop; welding and repair equipment; and telephone or radio communication systems.
Compressors are usually at the surface. They are usually of the two-stage type, and have an after-cooler as well as an intercooler, to avoid transporting any heat of compression into the heading, which is often too hot already.
Alternating current is used. When possible, it is purchased from a utility. It is usually stepped down to 220 or 110 volts at the entrance, but on some jobs is taken in at several thousand volts, in armored parkway three-wire cable. Dry transformers (oil-filled ones are a fire hazard underground) are set about 1,000 feet (305 m) back from the faces, and advanced in long jumps as progress warrants. This system avoids the power loss and voltage drop associated with long-distance transmission of low-voltage current.
There may be three electric circuits in the tunnel, a 220- or 440-volt for power, a 110-volt for light, and a high-voltage line for firing explosives. Some operators standardize on 220 for both lighting and power. Sometimes 220-watt bulbs are a nuisance to get in the United States, but they have the advantage of being useless in an ordinary lighting circuit, so they are seldom pilfered.
No drill dust can be tolerated. It may be suppressed by detergent or foam, or drowned in wet drills with water supplied through pipes from outside the tunnel.
Surveying. Tunnel sections meet each other far from their portals or shafts, sometimes after curves, with uncanny precision. Differences usually vary from a small fraction of an inch up to several inches (centimeters). These are too small to be noticed on the walls, but are measured at the surveyed centerline (axis).
An underground direction is obtained by establishing a baseline at the surface, running close to the line of the tunnel. This is very carefully done, and it is marked at frequent intervals by permanent monuments, with exact points pricked into metallic bolts embedded in concrete.
Two plumb bobs weighing 20 to 30 pounds (4.1 to 13.6 kg) each are suspended close to the bottom of the shaft by piano wire from the surface. They are as far apart as shaft width permits. Vibration and tendency to swing may be dampened by hanging them in pails of water. Very careful observations are taken of the wires at the surface, relative to the proposed tunnel centerline. Direction is identical with that of the same wires at the bottom.
Careful observations are taken of the bottom part of wires, using a very accurate instrument and special sighting devices. Readings are taken over and over, and the results averaged. The tunnel line is then established in the correct direction, by reference to surface readings.
This work must be done at a time when workers and equipment are not working, as ventilating currents and vibration can disturb the wires.
The line is extended through the tunnel by laser beam and/or transit, and marked on spads (markers) driven into holes drilled in the roof.
Exploration. Tunnels are seldom driven blind. Preliminary drilling is done along the route to determine the type of rock, the amount of water to be expected, and the danger of mud slides. Test holes are drilled from the surface, usually with diamond drills that can bring up cores for inspection.
Diamond drilling may also be done from the heading, where dangerous conditions are expected. This precaution has often revealed the presence of such quantities of water or unstable soil ahead, that disaster might have resulted had it been broken into by a full-face blast.
Figure 9.40 shows extensive core drilling that was done for a sewer tunnel under the East River, New York City, in order to find a way to avoid a dangerous seam of decayed rock.
Dangers. Underground work naturally is very dangerous, and it is greatly to the credit of tunnel workers and labor departments that there are so few accidents.
The most evident danger is that of collapse. Most soils and many rock formations will slump rather quickly into any hole cut under them. In any given material, this tendency increases markedly with depth. Below 500 feet (152 m) even apparently firm rock may creep, and break off slabs with explosive violence. There is always danger of loose pieces falling.
Caving and breaking off are combatted with compressed air, timbers, steel and concrete linings, and holding rock bolts.
If the soil will not stand at all without support, bracing must be installed ahead of the digging; or the heading protected by a movable shield.
Water, with or without accompanying soil, may break into a tunnel in such volume as to flood it completely within minutes. Escape of workers may be difficult, machinery is apt to be abandoned, and an expensive and tedious job of sealing off the water and pumping out the tunnel is often required before work can be resumed.
FIGURE 9.40 Underground exploration and tunnel detour. (By permission from “Practical Tunnel Driving” by Richardson & Mayo, McGraw-Hill.)
Fire must be carefully guarded against, particularly on jobs using compressed air and/or timbers.
Air conditions are difficult to keep healthy. Drills produce rock dust, and most air-powered machines have foul, oil-charged exhausts. Explosives produce fumes. Some clay and rock formations give off unpleasant or poisonous vapors. The increasing use of internal combustion engines underground makes tremendous demands on ventilation systems that try to clear out dangerous or irritating exhaust gases.
Silicosis, a lung disease caused by breathing dust from rock drills over long periods, was formerly one of the greatest health dangers of tunnel work. It now can be almost entirely avoided by wet drilling and good ventilation.
The Delaware Aqueduct used 10,000 feet (3,050 m) of air per minute at each heading. A standard estimate is 1,500 feet (457 m) per drill.
Shallow rocks are cool, but in deeper work an increase of 1°F (1°C) for every 50 feet (15.2 m) of depth can be expected.
Noise is deafening, particularly during drilling, as the sound echoes back and forth in the confined space. Diesel engines add to the uproar.
MSHA. The Department of Labor in the United States includes the Mine Safety and Health Administration (MSHA), which oversees and regulates the conditions that could be hazardous for workers in gassy metal and nonmetal mines. Although they were not originally intended to apply to tunnels for highways, railroads, water pipelines, and utility lines, the MSHA rules should be and are frequently applied in these tunnels.
MSHA defines a blowout as a sudden, violent, and unplanned release of gas or liquid due to reservoir pressure in petroleum mines and an outburst as a sudden, violent release of occluded gases and solids under high pressures from a geologic formation. Either a blowout or an outburst could happen in the types of tunnels discussed in this chapter. For the sake of reference, the mine atmosphere is tested at any point at least 12 inches (30.5 cm) away from the face, rib, back, and floor of the tunnel.
Requirements. The MSHA standards contain operational requirements that must be satisfied to meet the regulations for worker safety in the excavation. The standards stress tunnel ventilation to control airborne contaminants, including natural methane gas and equipment exhaust. There are minimum airflow quantities for ventilation. If air passing a particular input point has a mine atmosphere with more than 0.25 percent methane, it cannot be used for ventilation. Testing for methane must be done at least once every work shift. The volume of air ventilating each face at a working place shall be at least 20 feet per minute (6.1 meters per minute) multiplied by the open cross-sectional area in square feet (meters) of the entry. According to that minimum, if it is a 10-foot (3.05-m)-diameter tunnel, there needs to be at least 1,570 cubic feet per minute (20 × 3.14 × 25) [44.5 cu.m per min. (6.1 × 3.14 × 2.32)] of ventilation air flowing to the face. However, the required minimum at the face is 2,000 cubic feet per minute (56.6 cu.m/min).
The tunneling equipment is to be “permissible,” that is, it does not emit gaseous exhaust or electric sparks that could ignite a gaseous mixture. Therefore the best equipment to use is electrically or compressed-air powered. Within 100 feet (30.5 m) of the tunnel face or bench, the equipment used must be permissible, except that nonpermissible front-end loaders and haulage trucks equipped with methane monitors may be used at the face or bench after blasting. It is recommended that workers be removed from a blasting area for at least 30 minutes after a blast. MSHA requires that methane-monitoring devices be installed on continuous mining machines, and that should apply to tunnel-boring machines also.
Shafts—vertical passages between the tunnel and the ground surface over it—are required for the majority of tunnel jobs. They are sometimes the only access. Even when there are portals, shafts shorten the time required to do the work, as each makes it possible to work on two extra headings. In addition, underground hauling is a headache, and runs should be kept as short as practical. Some shafts are part of the permanent tunnel project.
The advantages of shafts must be balanced against the considerable expense of sinking and equipping them.
Shaft location may be chosen to keep depth to a minimum, as in the troughs of valleys over the tunnel; to take advantage of an easily worked or stable formation; or on the basis of surface conditions such as good access, nearby dumping areas, cheap land, or distance of shielding from populated areas.
Size. Shaft size is highly variable, depending largely on the volume of material it must handle and the size of objects that must be lifted and lowered. A minimum size, about 11 × 13 feet (3.35 × 3.96 m) inside the lining, accommodates a single hoist and supply elevator and a ladder-way. The ladderway includes a ladder, electric and high-pressure air lines, water pump discharge pipe, and ventilator ducts, all of which must be protected from swinging loads or falling chunks.
The headframe is a tower of prefabricated steel, as in Fig. 9.41, or may be built with timbers. This carries the hoist sheaves, dumping mechanism, and the discharge chute or hopper. The hoist engine and winch are ordinarily in a separate structure nearby.
FIGURE 9.41 Shaft and heading equipment.
Soil Excavation. Digging is started with a clamshell, which can dig soft soil unaided, and remove hard soil and rock after they have been loosened. One or two signalmen direct the operator’s movements, as he or she cannot see the bottom, and any wrong move with the heavy bucket might be disastrous to the workers. The clamshell is ordinarily not used below a 25-foot (7.6-m) depth.
The next stage may be to replace the digging bucket with a light bucket or container that is lowered to the floor, and loaded by hand or by equipment suited to the cramped work space. The container is raised out of the shaft by the hoist line, swung to the side, and dumped by a trip device or by hand. This may be used to a depth of 100 feet (30.5 m), or a direct transition may be made from the digging bucket to use of the headframe hoist.
A special small clamshell may operate from a platform close to the bottom, loading the containers that are lifted past it to the top by the main hoist.
Blasting. In shaft rock blasting, all the holes are tight—that is, there is no open face to permit sideward throw of the rock—so that close drilling and heavy loading are the rule. It is necessary that the rock be cut back cleanly to the digging lines and important that overbreak be kept to a minimum, because of the high expense of removing muck and the frequent requirement of filling all spaces outside the lining.
Figure 9.42 shows typical drilling patterns for shaft and tunnel work. A set of two or more converging angle holes (wedge holes) is drilled, and other sets of straight or slightly angled holes next to them, until the rim has been reached. The wedge holes are heavily loaded, so that they crush and kick out the rock between them, making an opening into which the rock around can move sideward when the next ring of holes is fired. These in turn make space for the next set. Firing is best done by short-period delays.
In (B) the floor is lowered on only one side in each shot. This permits drilling to be resumed on one side while muck is being loaded from the other.
Figure 9.43 shows a burn hole shot, with the center holes parallel instead of angled.
The blast is fired from the top, after all workers and equipment are out of the shaft, except that in very deep work some equipment might be merely raised far enough to be out of immediate danger.
After the explosion the bottom will be full of fumes, which would take a long time to dissipate naturally. These may be blown out by lowering the tool air lines with the ends open, or extending low-pressure ventilating ducts to the bottom. (These have to be dismantled or pulled back a considerable distance to avoid damage from the blast.) A suction line (foul air duct) is more effective at cleaning the air than a blow or pressure line, as the fumes tend to settle.
Some shafts are large enough to provide space to load the muck by machinery, but in many of them it is tossed, rolled, or hand-shoveled into buckets or skips, that are removed by the hoist when filled. The best fragmentation for this type of loading is usually one-person stone, that is, pieces that one worker can handle conveniently.
FIGURE 9.42 Wedge drilling in shaft.
Drilling can be resumed as soon as part of the bottom is cleared. Six- or 7-foot (1.83- or 2.13-m) steels giving a 5-foot (1.52-m) penetration are often used, but longer or shorter ones may be better in particular circumstances. Hand and wagon drills are standard, although special jumbos have given good results.
Working Up. When a large shaft is required as part of the finished job but is not needed for the early tunnel work, it may be more economically cut from below. With this method the blasted rock falls to the tunnel floor, and is removed through the portal.
The glory hole method is to sink a small pilot shaft to the tunnel, then to dig the large shaft from above, blasting or pushing the muck into the small shaft so that it will fall to the tunnel. See Fig. 9.44. Alternatively, a raise boring machine can enlarge a circular shaft to full diameter by operating through the pilot bore. In that case the pilot bore must be centered in the shaft’s diameter. See Chap. 20.
FIGURE 9.44 Raise and glory hole shaft excavation.
Shaft Lining. In soft ground the shaft is protected from caving by setting sheeting planks or sheet piling, held by whalers, in much the same way as described earlier for ditches and basements. The whalers are interlocked at the corners to hold each other in position, and additional divider beams may be run across between the ladderway and the hoistway. See Fig. 9.45.
If the soft soil is too deep to be held by sheeting driven from the top, successive layers can be driven from inside the shaft, inside the upper ones. If the sheeting is driven with an outward batter, shaft size will be preserved. If driven straight, the inside diameter of the shaft will decrease, so that the top would have to be oversize to allow it to be full size at the bottom.
Shaft lining or timbering is required for the hoist and to a lesser extent for the utilities, even when the soil or rock is self-supporting. When not needed for wall support, timbering may follow 20 or more feet (6.1 or more meters) behind the digging, to avoid interference with the work, and damage to the lower section in blasts.
FIGURE 9.45 Timbering for small shafts. (By permission from “Practical Tunnel Driving” by Richardson & Mayo, McGraw-Hill.)
The lining must be supported vertically to prevent it from slipping down. Each new set is fastened by hanging bolts to that above, and every 100 feet (30.5 m) or so horizontal notches or shelves are cut into the walls to provide fresh support.
Timbered shafts are usually rectangular, while metal or concrete linings call for circular cross section. Steel ribs are made up, curved to the proper arc, and divided into two or three pieces that are lowered endways and then supported by hanging bolts until fastened into a full circle. The actual lining or lagging may be sheet piling or similar material, or the ribs may be built as liner plates, with curved flanges which butt against those above and below to make a continuous sheet.
A continuous lining or lagging is used in soil that might squeeze between ribs or timbers, or in rock that scales or breaks off so that falling pieces would endanger workers. Unlagged walls may often be kept intact by spraying with shotcrete or a bituminous mixture.
Drainage. Most shafts are wet. If there is only a little water, it can be bailed into the bucket and hoisted with the muck. More often it is removed by a pump with a discharge line reaching to the surface or, if the height is great, to one or more pumps that help push the water out of the shaft. All pumps used in deep shaft work should be able to develop very high discharge pressures, so that a good lift can be obtained between boosters.
If water conditions are severe, the area may be predrained by sinking 4- to 12-inch (10.2- to 30.5-cm) holes with rotary drills, and pumping from them. Depth is too great for ordinary well point work from the surface, but in flowing ground well points may be sunk from the shaft bottom or sides, and the water rehandled by the regular pumps.
A deep, wet shaft should have gutters and sumps at intervals, to catch water running down the sides. Pumping to the top from intermediate points may be more efficient than allowing it to get down to the bottom and raising it from there.
Shaft Sinking Machines. A large and increasing percentage of new shafts are cut in a single or dual operation by gigantic drills, described in Chap. 20.
A heading is a digging face and its work area. Conventional tunnel driving is discussed here. Tunneling machines (borers or moles) are described in Chap. 20.
When the shaft has reached the level of the proposed tunnel floor, two headings are started, one in each direction along the line of the tunnel. In addition, the foot of the shaft may be greatly expanded for storage and maneuver space, and one or more rooms may be built to house compressors, pumps, and other plant equipment.
At first only a single set of tunnel-driving equipment may be used, as there will not be space enough for two, and greatest efficiency will be obtained by drilling at one face while mucking at the other.
Drilling patterns may be similar to those described for shafts—wedge or burn holes, and successive rings breaking into the crushed-out area. The whole face is usually drilled and blasted in one operation (full-face attack), but a small tunnel (drift) may be drilled full face, blasted, and cleaned out, then enlarged by radial drilling; or the top may be kept ahead of the bottom (bench-and-heading method). See Figs. 9.46 and 9.47.
Pilot Tunnel. Shafts may be partly or wholly replaced by a small pilot tunnel, driven parallel and close to the main tunnel. Crosscuts are driven from this to the main tunnel, wherever new headings are to be started. The main tunnel is opened up with a center drift, and enlargement started after it has been cut through enough that both tunnels can be used for traffic.
The extra tunnel may be used for ventilation, both during the work and afterward. It permits a great many operations to be performed at the same time, and may save considerable expense in sinking shafts. This method has been used chiefly for long railroad tunnels through mountains where depth was too great for shafts.
FIGURE 9.46 Tunneling with big equipment. (Courtesy of Atlas Copco.)
Drilling. The standard tool for small tunnel drilling has been the drifter, a medium-weight hand drill with a hand or automatic feed, mounted on a vertical column or a horizontal bar of such length that it can be secured between the floor and roof, or between the sides, by screw-jack ends. Because of the weight of the columns, they become impractical for full-face work in tunnels of greater cross section than 10 × 10 feet (3.05 × 3.05 m).
The drifter permits the drill crew to resume work on the top of the face as soon as blast fumes have cleared away, with the drill operators standing on the pile of muck until it is dug away. They can drill the bottom after it has been cleared.
Larger tunnels may be done by the heading-and-bench method shown in Fig. 9.47. This permits the use of drifters on short columns for the advance, and approximately vertical jackhammer or wagon drilling for the bench. Sometimes the heading is extended far ahead of the bench, and has its own hauling equipment that dumps over the bench face into other cars, or into a pile to be dug away.
For larger tunnels to be drilled and blasted, the standard method is to use a drill carriage (jumbo) on which power feed drills can be mounted so as to reach all parts of the face at correct angle and to correct depth. Each drill usually does several holes. It can be positioned by hand, or by mechanical, air, or hydraulic controls. Such jumbos may be so constructed as to straddle hauling equipment, so that it need not interfere with removal of muck. They may also carry a cherry picker crane to pick up empty cars to switch loaded ones through. They are backed away from the face before each blast.
On very large tunnels jumbos may be used on both levels of heading-and-bench work.
Usual drilling depth is 10 to 12 feet (3.05 to 3.66 m), but in any case is seldom deeper than two-thirds the smallest dimension of the tunnel.
Figure 9.48 shows the typical full-face drilling patterns.
FIGURE 9.48 Drilling pattern, large tunnel.
Bits. Recently tunnel drilling has been partly standardized to use steels threaded to carry detachable bits. These may be multiuse types that can be sharpened by grinding, or sharpened and reshaped by hot milling; one-use or throwaway bits that are discarded when dull; and carbide insert bits. The carbide insert bit has caused a spectacular advance in speed and ease of hardrock tunneling. Carbide outwears steel at an average of about 100 to 1, and gives much more rapid hard rock penetration. The time of handling, transporting, and processing bits is reduced from a major to a minor problem.
Loading. Water-resistant explosives with good fume characteristics are desirable in underground work. These qualities are found in gelatin dynamites.
When all holes in a face have been drilled, each is blown out with a high-pressure air jet to remove loose cuttings and water. Cartridges are slit (unless the explosive has been damaged by water and the hole is wet) and tamped firmly with a wooden pole. It is common practice to place the primer after the first cartridge, with the cap pointed toward the collar of the hole.
Stemming may be taken from the drill cuttings. It is most convenient to use if wrapped in paper bags of the same size as the cartridges. If this material is very high in silica, its use as stemming might increase the silica in the air enough that prewrapped blanks supplied by powder manufacturers might be preferred. There are also wood and rubber plugs that are very satisfactory.
It is good practice to place a wad of paper between the explosive and the stemming, so that the powder can be easily and safely located in case of a misfire.
There is danger of premature explosion from stray currents. A common precaution is to take down or “kill” all electric wiring within 500 feet (152 m) of the face before starting to load. Safety flashlights, of hand or cap models, or headlights from a battery locomotive can be used. It is sometimes a question whether the poor lighting obtained offers as much of a hazard as the electricity would.
Even the complete absence of electricity on the job would not guarantee a tunnel face against currents, as underground water is often highly mineralized and will conduct a charge for long distances. Metallic ores may be excellent conductors.
The precautions described earlier for blasting in the presence of electrical hazards should be followed.
Firing. Any wiring hookup can be used—series, parallel, or parallel series, depending on the preference of the blaster. If 440-volt electricity is available, it is preferred for firing, although 220 or even 110 will do. Regular blasting machines are also used, but they should not be kept in the tunnel when not in use, because of possible damage from dampness.
All equipment is moved 500 to 1,000 feet (152 to 305 m) back from the face, as rocks caroming off the walls can travel long distances. Compressor pipe can be left fairly close to the blast, but ventilation conduit must be stripped way back.
Checking. It is important that a thorough check be made after the blast for misfires. Tunnel work brings a large number of workers into close contact with the heading, and any accidental explosion during mucking or drilling would be disastrous. The best check is inspection by experienced people.
If an unexploded hole is found, and the wires are intact, they can be hooked up and fired. If the wires are missing, the stemming can be washed out by a water jet, and a new primer inserted and fired. Or a parallel hole, about 2 feet (0.61 m) away, can be drilled, loaded, and fired. The muck must be inspected for unexploded cartridges.
Loading. In small tunnels, blasted rock may be dug by hand, although the excellent mechanical loaders adapted to work in tight quarters that are now available, and the rising price of labor, are steadily reducing the practice. Output for the loading gang is generally figured at about ½ to ⅔ yard (0.38 to 0.51 cu.m) per hour per person, although one person may load up to 2 yards (1.5 cu.m) under favorable conditions. The difference lies in the work of loosening, and of handling cars, and in other delays.
The swell or “growth” of rock in passing from the solid to the blasted state averages about 50 percent. In tunnels, mucking is usually calculated in terms of loose yards, in mines in number of tons loaded.
Slick sheets should be used in connection with hand loading. These are thin steel plate, ¼ or 
inch (0.64 or 0.79 cm), in pieces about 4 × 6 feet (1.2 × 1.8 m), with holes punched for convenience in picking up for moving. They are laid to cover the tunnel floor for 10 to 25 feet (3 to 7.6 m) back from the face before each shot. Large rocks are picked up and thrown into the cars individually, while the finer material is dug by shovels that slide easily along the metal surface.
Mechanical loaders include full-revolving shovels with short booms and proportionately larger buckets, that move on either crawlers or rails.
Special tunnel-mucking machines are available in large variety. Most are rail-mounted, although crawlers are gaining in popularity. The bucket can be swung from side to side to reach the full floor area, and is filled by pushing into the pile.
It is then lifted, in some models over the machine to discharge into a car or conveyor belt behind; in others it loads a built-in conveyor that discharges to the rear. In either case, the car may be coupled to the mucker so that it is always in loading position.
One method of tunneling in hard rock, 35,000 to 50,000 psi (2,400 to 3,450 bars) strength, is to drill a pilot hole, perhaps 5 inches (127 mm) in diameter, and then ream it to a larger size, maybe in several increasing size steps. This is what was done to extend water and sewer lines in Atlanta, Georgia, USA, where they were tunneling through granitic rock. The lines up to 26 in. (660 mm) in diameter were reached in several reaming passes.
Another approach is to use what is known as a Small Boring Unit (SBU), which is suitable in rock with strengths of 4,000 to 25,000 psi (276 to 1,724 bars) and for distances up to 500 ft (152 m). The SBUs are used in bores from 24- to 72-inches (0.61 to 1.83 m) in diameter. It is launched in the started pit, with its cutting head welded to the lead casing. Then the SBU applies its thrust and torque for the cutting head through the lead casing. When it has progressed enough to add new casing, the starting weld is broken and the SBU works off of the new casing.
A tunnel may be cut to full size in a single operation by boring with a tunneling machine, sometimes called a mole. These machines are described in Chapter 20. See Fig. 9.49.
A tunnel borer grinds, chips, or digs its way through formations, by rotary or oscillating motion of cutter teeth, and deposits the muck onto a conveyor belt for discharge into haulers at the rear.
In addition, it may and often must provide for placing steel or concrete linings around or behind itself. Such lining can be used to take the thrust of its crowding force on the face.
These machines can usually be disassembled sufficiently to be brought down a large shaft, and assembled in a conventionally dug tunnel section at its foot. However, whenever possible, they start work at a portal (outside entrance).
Almost any type of hauling unit may be used in a tunnel, from a wheelbarrow to an off-the-road ten-wheeler. It is a matter of tunnel size, speed of driving, ventilation, and preferences of the management.
The traditional system is small muck cars pulled along narrow-gauge tracks by electric locomotives. The locomotives can take power from either batteries or high lines, and start at 4 tons (3,630 kg). There is an increasing use of diesel locomotives with exhaust conditioners in well-ventilated tunnels. See Fig. 9.50.
FIGURE 9.49 Bored tunnel face.
Cars are usually side-dump types, although many special constructions are found. The width is governed by the tunnel and the gauge of the track, and should be small enough to allow passing in the tunnel. Car width is generally about twice the track width.
The capacity of the car may be limited by switching arrangements. If they are pushed by hand, capacity is limited to 1 or 2 yards (0.76 or 1.53 cu.m), as heavier cars will need to be pried along the tracks, rather than shoulder-pushed. The car must be low enough to go under the discharge of the mucking machine being used. If hand-loaded, it must not be over 4 feet (1.22 m) high.
The loaded muck cars are hauled to the shaft and run into hoisting cages, in which they are lifted to the top, where they are dumped by side tipping. There are also special cars that can be lifted directly, without entering a cage. Or they may be dumped at the bottom into a hoisting skip.
The perpetual problem in tunnel haulage, which becomes more acute as size decreases, is bypassing the empty cars (or trucks) going to the face around the full ones coming away from it. Empty cars may be switched to the side; or if they are small, lifted or pushed off the track by hand, where there is space for only one track. Larger ones may be handled by a cherry picker. In either case the spotting arrangement shown in Fig. 9.51 may be used.
The locomotive pulls a string of empties into the heading and stops to let the cherry picker take up the rearmost car and set it aside. The locomotive then backs far enough that the car can be replaced on the track in front of it; then pushes that car up the loader. While it is being loaded, it backs so that another car can be picked off.
When the car is loaded, the locomotive couples to it and backs past the cherry picker, which places the empty in front of it to be pushed to the face. While it is loaded, the rear empty is again set aside, to be pushed in on the next cycle. When all the cars are filled in this manner, the locomotive pulls them to the shaft.
FIGURE 9.50 Diesel-powered mine locomotive.
In a tunnel of sufficient height, a movable framework called a Grasshopper, Fig. 9.52, can be used. This allows the empty cars to be moved over the loaded ones, and can be pulled up to the face by the loader.
A conveyor belt may be set up so that a full train of cars can be backed under it, and loaded one by one from the front to back.
FIGURE 9.52 Grasshopper overhead switch.
Conveyor belts can also be set up to haul from the face to the shaft. No switching arrangements are required, but this unit cannot be used readily to bring supplies from the shaft to the face; considerable work is involved in dismantling or protecting it for a blast, and there is constant work adding sections to keep it in touch with the digging.
Diesel-powered trucks are used for large tunnels. They carry much bigger loads than mine cars, and if sufficient width is available to make passing possible, they get past each other with fewer complications than rail-mounted carriers. The shuttle types, such as the Dumptor, which are equally comfortable going backward or forward, are often better adapted to the work than those which have to be turned in the tunnel.
The use of internal-combustion engines fouls the air, so that very good ventilation is required.
Exhaust Gas. The exhaust from a gasoline engine contains carbon monoxide, an odorless but poisonous gas that soon makes any closed-in place deadly to life. Amounts of carbon monoxide that are not sufficient to cause unconsciousness or death may temporarily damage judgment and reasoning power, causing an increased danger of accidents.
Diesel exhaust contains little carbon monoxide, but it is rich in various chemicals that smell badly, are irritating to eyes and throat, and that fog up the air so that visibility is dangerously reduced. This last difficulty is increased by the usually bad lighting in a tunnel.
The danger from gasoline engine exhaust has largely prevented use of this type of power underground. Diesels are used in spite of the irritation and danger they cause. Their presence is partly compensated for by increasing the ventilation, but conditions do become very bad. They are often made worse by an astonishing lack of care in adjustment of the engines. Diesel trucks sometimes emerge from tunnels belching black smoke, presumably caused by defective or souped-up injectors.
Various types of scrubbers using water and chemicals to dissolve and neutralize gases, and secondary catalytic oxidizers that serve also as mufflers, are used to make internal-combustion engines acceptable underground. These are described in Chap. 12, under Exhaust Conditioners.
Good ventilation and lots of it are a basic requirement, even when such devices are efficient. The most they can do is reduce the exhaust to carbon dioxide and water. Carbon dioxide is not poisonous or irritating, but in sufficient concentration it has a suffocating effect that can cause impairment of judgment, unconsciousness, and death.
Groundwater is a problem in most tunnels, and may be the principal one in some. Many mining tunnels, some of them miles in length, are made solely to lower the water table. There may be seepage all along the line, adding up to a considerable volume to be drained or more often pumped away. Gushing springs may be exposed by any blast, or may open up from seepage points well behind the face. Underground lakes or rivers may be encountered that are capable of flooding the work in spite of continuous pumping. Veins of soft, water-soaked soil may be found in hard rock, that may break into and fill the tunnel.
The first necessity is adequate pump capacity. The tendency is to underestimate requirements, largely because pumps and lines are expensive, partly because even careful exploration from the top seldom reveals the full quantity and pressure of water that may be encountered.
If a tunnel runs uphill from a portal, drainage may be by natural flow through a ditch cut along the side. If an upgrade from a shaft, it can be drained to a pump inlet at the shaft foot. This arrangement is easy and inexpensive but seldom satisfactory, because of repeated blocking of the ditch by rockfalls from walls or from hauling equipment, resulting in water running over the floor, making it sloppy and often undermining the track or spoiling the road surface. The ditch also takes up more space than a pipe, and there has not yet been a tunnel with floor space to spare.
The conventional arrangement is to pump all water. A small centrifugal pump, usually air-driven, is kept near the face, and takes from a sump and discharges into a pipe running back toward the portal or shaft. Another sump is provided every 500 to 1,500 feet (152 to 457 cm) back to collect local water for another centrifugal, usually electric-powered pump. Each pump may discharge into the sump behind it, which is kept down by another pump, usually of a larger size. Another arrangement is to have all pumps discharge through check valves into a common discharge line. A powerful electric pump of the piston or centrifugal jetting type is installed at the shaft bottom, and as many boosters as are required for the lift installed at intervals in niches in the shaft.
Pipelines vary from 1½ to 10 inches (3.8 to 25.4 cm) in diameter.
The pump or pumps at the base of the shaft are sometimes placed in a sealed room, with power and control directly from the shaft top. In other cases the pumps are in the open, but are of the submersible type. These arrangements permit use of the units along with emergency pumps if the tunnel should be flooded.
Grouting. Water inflow can often be checked by grouting. This may be done by drilling deep into the rock in the direction of the supposed source of the water, sealing in pipes with cement, and then pumping cement and water grout through them, either straight for seepage or mixed with sawdust or shavings for gushing flow. This may be done in advance of the tunnel driving in very wet areas, by fanning the grout holes out from the face and edges of the heading, as in Fig. 9.53.
Grouting is also done through completed linings, either to check water or to fill in spaces between it and the wall. Grout pipes may be cemented into a concrete lining when it is poured.
FIGURE 9.53 Exploration and grouting holes.
Successful grouting of a wet seam sometimes merely diverts the water so that it enters the tunnel at another point that was previously dry. This also may be grouted, but a point may be reached where the contractor either installs a complete concrete lining or gives up the effort to seal off and relies on pumps.
The aboveground uses of grouting were discussed in Chap. 6.
Ground pressure in rock tunnels is difficult or impossible to estimate. In firm formations there will be little or no pressure until depths of greater than 500 feet (152 m) are reached.
However, there are soft, joined, or laminated formations that will scale off or fall from a flat or moderately curved roof, until a Gothic or pointed arch develops, after which it will be self-supporting. If bracing is done only to support the roof, it is a question whether it will be more economical to cut up to a stable roof line, and avoid placing of supports. See Fig. 9.54.
In any roof problem, width is a very important factor, as wide spans will drop pieces or fall in much more readily than narrow ones.
Many rock tunnels are perfectly safe without any bracing. Others get by without accidents. But very often is is necessary to place supports directly after the digging, or within a few days. Also, the majority of tunnels outside of mines are more or less permanent in nature, and except in very firm rock, will require lining to prevent deterioration and to reduce or eliminate maintenance.
Support or lining may be wood timber, steel ribs, plates or bolts, or concrete. Concrete is frequently placed inside one of the other types of support.
Timber is the oldest material used, and is found in ancient tunnels. Concrete was used to some extent by the Romans, and has become the standard for permanent installations. Steel liners and roof bolts are quite modern developments, and are rapidly replacing timbering.
Timbering. Figures 9.55 to 9.57 show some designs for timbering. The square-set framing is confined to small tunnels, and various forms of arch construction can be used in quite large ones. The arch may be supported on posts supported in the floor, or rest on a springline shelf (hitch) cut in the sidewalls. Support may vary between these methods with changes in ground, or in shape of the edges.
FIGURE 9.54 Cutting crown for self-support.
FIGURE 9.55 Timber set, small tunnel.
FIGURE 9.56 Timber arch on posts.
FIGURE 9.57 Timber arch on hitch.
Posts should be fastened to the wall plate by dowels, lag bolts, or scabs (nailed-on pieces) so that they cannot fall if relieved of weight.
The weight of timbering varies with expected ground pressure. Sometimes it is merely a light roof to catch light rockfalls, at other times a high-strength lining designed to resist squeeze from all directions, including the bottom.
Where timbering ends at a portal, or at an enlarged shaft base, it must be securely braced by diagonal beams, as in Fig. 9.58, so that any compression developing in the tunnel will not squeeze it out.
Packing. In rock or soil that tends to push in, it is important not to leave any space between the lagging and the wall or roof, as any inward movement will increase the instability of the ground, and may cause it to exert tremendously more pressure than if it had been held in its original position.
An exception is found in swelling or squeezing ground that is allowed a limited space for movement.
Initial movement is prevented by packing the space between the lagging and the rock. The most economical system is to use a dry packing of fine muck, which is shoveled behind the planks as they are placed. At the crown it must be thrown in from the end and securely rammed—a tedious, disagreeable job that is seldom well done.
Dry packing may also be done with pea or birds-eye gravel, shot into place with pneumatic guns either through holes in the lagging or from the end. Its use is more common in soil than in rock.
FIGURE 9.58 Timber bracing at portal.
Lean concrete, with a cement–sand–small stone mix of 1:3:6 or 1:4:8, can be shot into the arch with a pneumatic placing tool. This must be very dry so that it will not leak through cracks between the planks. This is done after the set has been erected and securely blocked, as the fresh concrete may impose very heavy stress.
Lagging is usually set closely (skintight) in the crown. On the walls it may be widely spaced or lacking, as even if the rock squeezes in, the spans between timbers are too short to allow bulging. Under very heavy conditions the timbers may be set skintight, so that lagging is not needed.
If the tunnel is to be concreted, the lagging may be placed inside the timbers, to provide a smoother outer form that saves concrete yardage. The disadvantages are that much more packing is needed and that the fastening is under tension rather than compression, so that heavy pressure may make the lagging pop off the timbers. The effect may be cumulative, as yielding of one fastening increases the strain on the next, so that a considerable length may give way at one time.
Steel Ribs. Steel supports are standard in tunnel work. They are easier to handle, and allow substantial saving in excavation. This is because for a given strength, they are only half as thick; and the projections of ribs into a concrete lining are counted as reinforcing. In timber construction, the outside line of the concrete is figured as the inside line of the timbers, and the concrete used in to fill out to the lagging is largely figured as waste. On small tunnels the saving by use of steel in excavation may be 30 percent and in concrete, 50.
FIGURE 9.59 Roof bolts in a tunnel arch.
However, steel liners are more vulnerable to blasting damage, and do not give warning of impending collapse under load by groaning, as timbers do.
The steel ribs are made in two pieces, occasionally more. They are brought in endways and set up individually. The lagging may be wood planks or steel liner plates. If the former, the ribs must be well strutted to each other to keep them in line.
As in the case of wood, steel lining may be only a roof or crown support based on shelves at the spring line in the sidewalls, or a complete tunnel enclosure.
Roof Bolts. It has been found in mining and tunneling operations that unsafe rock will often support itself safely over wide spans if it is reinforced with steel bolts. (See Fig. 9.59.)
In laminated (thin-bedded) formations, the effect is similar to that obtained in plywood and other layered-wood constructions. Several weak and thin layers may be very strong when bonded together. In jointed and fissured rock, the bolts, if used properly and in sufficient numbers, restore to the rock the massive strength it had before it separated into blocks and pieces.
Expansion bolts are used, rather similar to those that fasten wood framing to masonry. The type shown in Fig. 9.60 is made in ⅝-, ¾-, and ⅞-inch (1.59-, 1.91-, and 2.22-cm) diameter. The ¾ (1.91 cm) has a minimum breaking load of 22,500 pounds (10,200 kg) in regular strength, and 32,000 (14,500) in high-strength steel. Lengths are 2 to 8 feet (0.61 to 2.44 m), in 6-inch (0.15-m) steps.
FIGURE 9.60 Square-head mine roof bolt.
The bolt threads into the plug of an expansion shell that fits into a 1⅜-inch (3.49-cm) diameter hole. Ears on the bolt prevent it from sliding too far into the shell, so that tightening pulls the plug down into the shell, expanding it against the sides of the hole. (See Fig. 9.61.)
Roof ties may be used to support the roof between bolts. Wire mesh can be used in addition to the ties, or instead of them, where the problem is separation of small pieces.
Flat or dished (reinforced) plates of ⅜- or ¼-inch (0.95 or 0.64 cm) thickness and 6-inch (15.2-cm) diameter are used where ties are not needed. They are usually made with a 1⅜-inch (3.5-cm) hole. A hard steel washer prevents the bolt head from pulling through.
The drilled hole must be at least as deep as the bolt is long, and may be deeper. The bolt is usually provided with a hard steel washer and assembled to the shell at the factory. The shell is pushed through a plate and as far as it will go into the hole, and the bolt head is tightened with an impact air wrench, usually to a torque of 150 to 200 pounds-feet (20.7 to 27.6 kg-m).
The bolt head is held outside by the plate and washer, so the threaded plug is pulled outward as it is tightened, squeezing the shell against and into the walls of the hole. The grip of the shell in hard shale, sandstone, or limestone is usually greater than the breaking load of the bolt. In soft shale, grip is usually less than bolt strength. In rotten rock, grip may not be adequate.
Holding power can often be increased by injecting cement grout into the space around the bolt, after tightening.
A very strong grip is obtained by anchoring with resin. A capsule containing resin, fine aggregate, and a tube of hardener is placed in the hole, then broken and mixed by turning a bolt into it. After the mixture has set for 30 to 60 minutes, a bearing plate and hard washer are tightened onto the bolt with a nut and a torque wrench.
The washers offer the advantage of producing greater bolt tension with the same effort, or the same tension with less effort. The more uniform tension provides greater security.
Bolting requires from one-fifth to one-tenth the steel required for ribs and lagging, and under many conditions is equally strong. In addition, it saves the need of excavating space in which to set the steel structure, and reduces the amount of concrete required for permanent lining.
FIGURE 9.61 Installation of headed and slotted bolts.
Elimination of all ribs and timbering makes a tunnel easier to work in, as there are fewer obstructions, and it provides for a smoother flow of ventilating air.
Another important advantage is that the economy of the work causes it to be done on roofs that might be judged to be self-supporting if bracing were time-consuming and expensive. The bolts can also be installed right up to the face immediately after blasting, so that protection is available to the heading crew. As a result, their use in the rather wide range of conditions where they are applicable results in a marked decrease in roof-fall accidents.
Heavy wire mesh may be used to prevent falling of small fragments in between the bolts. In some instances gunnite is used to minimize air-slacking and spalling.
Rock anchor bolts, which are similar to the slotted mine roof bolts, are used along highway and railroad cuts to prevent rock falls and slides.
Concrete Lining. Installation of concrete lining is construction rather than excavation work (however necessary it may be to the excavation) and will be only briefly considered.
There are two general procedures—soft ground technique, in which it is placed immediately behind the digging and is necessary to the driving of the tunnel, and hard ground. Under the second heading comes work in rock that is self-supporting, and requires lining for permanence, scaling protection, or waterproofing; and unstable soil or rock that is adequately held in place by timber or steel.
The soft ground technique is to follow the heading closely, with some resulting interference between operations. Perhaps the most serious is maintaining a track for muck cars through the pouring operation, and across the freshly laid invert (floor). Steel beam bridges may be used to carry the track in this section.
The invert may be laid about 1½ inches (3.8 cm) low, protected with planking, and brought up to grade with a top dressing of thin concrete as a finishing operation after the tunnel is complete.
Traveling forms of various types are used. For fast schedules, it is essential to have telescoping forms that can be folded up and moved through other forms supporting more recently poured sections. On other jobs, forms are used that can be collapsed just enough to break away from the concrete surface so as to be moved ahead to the next section. In either case the forms are carried on carriages that may move on steel wheels and tracks or on rubber tires, depending largely on the muck haulage method used.
Breakthroughs. Sometimes, in spite of precautions, there will be a sudden rush of water or mud into the tunnel. This most often occurs at the face immediately after a blast. Sometimes the source is a limited underground pocket which will give no trouble after it has once drained off. At other times a stream or large body of water will keep up a continuous flow. If the water is muddy, or the flow is partly or wholly mud, an unstable soil formation has been reached which may give increasing rather than diminishing trouble.
In any case the first step is to seal off the face with a bulkhead (wall) as quickly as possible. Timber, sandbags, or sandbags with timber may be used. Occasionally timber may be backed with concrete.
The bulkhead must not be used as a dam while being built. Pipes should be built into it large enough to take the water flow until the structure is complete. Otherwise water pressure will tend to destroy the bulkhead as it is being erected, and conditions will be very dangerous to personnel. With water discharged through pipes, the structure can be properly and strongly made and keyed into the tunnel rim. The water can then be controlled by valves on the pipes.
The bulkhead should also be fitted with pipes for grouting and concrete placement. After the water has been shut off, grout can be injected into the space between the bulkhead and the break, and will sometimes work back along the water seam and stop or reduce the flow. Grouting may also be done through exploration holes drilled through the bulkhead and into the rock beyond. Over 90,000 bags of cement have been used to control one water pocket.
Further tunneling through such a spot is first in the form of drifts (small tunnels) each of which serves as a base for further grouting, until the ground is consolidated enough to drive the big tunnel.
Tunneling machines, as described in Chap. 20, have an important effect on methods and costs of solid rock tunneling.
Soft ground is divided roughly into the following subclasses, description of which is abbreviated from Practical Tunnel Driving by Richardson and Mayo, McGraw-Hill.
Running ground: Must be instantly supported. May be dry sand or gravel, quicksand, silts, and muds.
Soft ground: Roof must be instantly supported, but walls will stand vertically for a few minutes.
Firm ground: Roof will stay up unsupported for a few minutes, and the sidewalls and face for an hour.
Self-supporting ground: Will stand unsupported while the entire tunnel is driven a few feet (meters) ahead of the timbering.
The standard methods of driving through soft ground are forepoling with wood or steel, or working in a shield. The plenum method is keeping out soil and water with air pressure, with either forepoling or shield.
Forepoling. The use of plank forepoles was formerly the standard method of driving a tunnel through soft ground. While this technique has been largely replaced by steel liner and poling plates, it is still widely used on jobs too small to justify obtaining steel.
In forepoling, the tunnel is protected by timbering, and by breast boards set against the face. Planks are driven through slots cut in the breast board and supported cantilever fashion to make a temporary roof, under which dirt can be dug and permanent supports installed.
Figure 9.62 indicates the terminology of the parts and something of the method.
FIGURE 9.62 Forepoling soft ground.
Starting from a shaft lined with plank sheeting, a bent (cap or roof timber and two post supports) is set and securely braced close to the sheeting. Close-set holes are drilled through the sheeting in a double line just above the cap and a single line about 18 inches (0.46 m) below it. Double vertical lines are drilled just outside the posts (A).
A set of light forepoles or spiles are made of 2 × 6 (5.1 × 15.2 cm) planks 5 to 6 feet (1.52 to 1.83 m) long, sharpened to a chisel point on one end. A piece of sheeting is knocked out between the lines of drill holes above the cap, and a forepole is rested on the cap and driven through the hole, at an upward slant of about 2 inches (5.1 cm) per foot, for about half its length (B). Another bit of sheeting is cut or knocked out, and another forepole driven in the same manner, parallel to and touching the first. This process is repeated until the full width of the cap has been covered.
Spiles are then driven into the sides, flaring out about 2 inches (5.1 cm) per foot (meter), to a penetration 6 or 8 inches (15.2 or 20.3 cm) deeper than the roof pieces. These may be driven horizontally, or at an upward slant to keep contact with the roof.
A timber is now placed across the shaft, immediately above the free ends of the roof forepoles, which are then forced downward slightly by driving wedges under the timber. The poles are now supported on the cap and held down tightly by the timber and wedges at the rear, so that the front is supported cantilever fashion.
The sheeting is then broken out from the spiles down to the lower line of holes, and the ground allowed to run into the tunnel until it assumes its natural angle (C). The resulting slope will normally not extend back to the points of the forepoles, but will end at some intermediate position.
Next a horsehead, or false set, is placed under the poles about 2 feet (0.6 m) beyond the sheeting. This consists of a cross piece under the spiles, and a center post set on a small supporting block in the dirt (D). The spiles are then driven to their full penetration, substituting the support of the horsehead for that of the cap rear timber.
Earth is then raked in until the points of the spiles are almost uncovered. A board the width of the cut and about 18 inches (0.5 m) high is set vertically immediately under them. This serves as a breast board to keep more dirt from flowing in, and supports the spiles (E).
A cap timber is then set to line and grade, and is temporarily supported by a single center post. A “bridge” of 2 × 6 (5.1 × 15.2 cm) planks is fastened to the top of the cap but separated from it by 4-inch (10.2-cm) blocks (F).
The remainder of the side spiles are now driven. Some of these are tapered and are used wide end forward, so as to reverse the upward slope of the roof spiles and the upper few wall spiles.
The forward cap is now supported by a pair of beams resting on short temporary posts and wedged down from a cross timber. The remainder of the sheeting below the first cap is now broken out from the top down, and the dirt pulled into the tunnel and hauled away. Additional breast boards are set under each other as space becomes available, and held in place by cleats nailed to the side spiles.
When the floor is cut to grade, side posts (legs) are set on below-grade blocks and wedged up until they take the weight of the cap. Wedges are driven between the posts and the side spiles to tighten them. Sometimes a trench jack must be used to force the side spiles out while setting the legs.
The next set of roof spiles is entered through the bridge slot on top of the second cap, and is driven at the same upward angle. Space for side spiles outside the legs is obtained by knocking out the wedges as the spiles are placed for driving.
All spiles should be driven skintight (touching throughout their length) except at the corners, where 1-inch (2.5-cm) boards (lacing) are tacked on. When necessary, cracks are stuffed with excelsior, salt hay, or other packing to prevent inward leakage of soil.
Each timbering set must be braced securely to that behind it, as any shifting will severely weaken the structure. Spiles are usually driven with a sledgehammer or air hammer. Sometimes they are jacked in—a very tedious job—to avoid jarring the soil.
If the tunnel floor tends to get muddy, it should be floored, for convenience of workers and to avoid possible shifting or settling of the foot blocks. Sometimes floor spiles are driven if the bottom tends to boil up, but compressed air is a better way to combat this and other difficulties with excessively soft ground.
This method is relatively easy to follow in many soils, but it takes an experienced crew to get through boulders, flowing mud, and other difficult conditions.
Forepoles may also be used with steel ribs instead of timber sets.
The standard soft ground tunneling hand tool is a short-handled, round-pointed shovel, aided when necessary by a grub hoe (mattock), pickaxe, or crowbar, and often by paving breakers. Special grub hoes have one hammer face for use in driving wedges. In soft clay a curved two-handled draw knife can be used to advantage. It is pulled by two workers or a power winch, and slices the clay off in strips.
Liner Plates. Corrugated steel liner plates, curved to match the tunnel rim and supplied with drilled bolt holes in flanges or overlaps for fastening to each other, are increasingly used for soft ground tunneling. They are made in various sizes, with 16 × 36 inches (0.41 × 0.91 m) in common use. A plate of this size made of ⅛-inch (3.2-mm) metal weighs about 27 pounds (12.3 kg), and if ¼-inch (6.4-mm) stock, 53 pounds (24.1 kg). Short plates are available for fitting into the tunnel circumference.
Stiffening ribs are used when the tunnel is over 10 feet (3 m) in diameter, and for heavy loads in any size opening. They are generally not used when the same strength can be supplied by a heavier-gauge plate.
Liner plates are usually a temporary support to hold the tunnel until a concrete lining has been installed, usually a matter of hours or a few days after the digging. They are sometimes “robbed” for reuse immediately before the concrete is placed. The safety of this practice depends on the character of the soil, which is a matter for engineers to pass on in each case.
Liner plates are placed from the top down. A small section is dug ahead, the center plate placed and braced with a post or jack; and then the sides dug away to place the adjoining plates. These are supported radially on cleated center blocks, as in Fig. 9.63. When the spring line or base of the roof arch is reached, two 2 × 8 (5.1 × 20.3 cm) planks, called footing boards, are placed on each side. Wedges are driven between the two boards until they lift the arch of liner plates enough to take the weight off the jacks. The lower plates are then nailed to the boards to prevent slipping off them.
If the ground is too soft for this method, interlocked poling plates (Fig. 9.64), can be placed outside and forward of the completed liner, and jacked forward from inside.
Shields. Shields have become the standard equipment for driving major tunnels in soft ground. A schematic view of one is shown in Figs. 9.65 and 9.66. It resembles a tin can with an open back and controlled openings in the front. The front may be open, with grooves to allow setting a breast board or plates if necessary, or closed by a bulkhead with controlled ports. The back or tail is large enough to permit placing the tunnel lining inside it.
The shield is forced forward into the dirt by jacks based on tunnel lining. Doors in the front are opened to allow soil to flow in, or to be shoveled. In very soft ground where bulging of the surface will cause no damage (as under rivers or swamps), no dirt need to be taken into the tunnel, as it will be pushed aside by the pressure of the shield.
FIGURE 9.63 Setting liner plates. (By permission from “Practical Tunnel Driving” by Richardson & Mayo, McGraw-Hill.)
FIGURE 9.64 Steel poling plate and jack. (By permission from “Practical Tunnel Driving” by Richardson & Mayo, McGraw-Hill.)
A primary lining, which is most often of bolted cast-iron segments but sometimes of cast steel (for unusual stress), fabricated plates, concrete blocks, or timber, is constructed in the tail, which is long enough to protect a complete segment. This lining must be strong enough to not only resist full soil pressure, but also to take the thrust of the jacks that move the shield forward.
The outside diameter of the shield tail must of course be larger than that of the lining built inside it. A few plastic soils can be manipulated so as to close in smoothly on the lining as the tail moves away from it, but under most conditions the space must be filled. Failure to do so will leave the lining without proper side support, so that the arch will tend to sag.
Grout was originally the standard filler for this space. Grout plug holes were built into the liner pieces. When the tail cleared them, grout was forced into the bottom hole, with the next above used as an air vent. When grout appeared at the upper hole, the grout hose was transferred there, the bottom plugged, and injection continued. The full circumference was worked in this manner from the bottom up.
FIGURE 9.65 Tunnel-driving shield.
FIGURE 9.66 Shield in mixed face heading.
The amount of cement used makes this operation costly, and in addition the grout has a tendency to move forward along the outside of the lining and flow under the tail into the shield. Also, grout may work up to the surface, cause heaving of pavements, or break into sewers or conduits.
Gravel filling is used to avoid these difficulties. Bird’s-eye gravel (uniform size, passing ¼-inch (0.64-cm) screen, 33 percent voids) or similar sizes of slag or screenings can be blown by an air gun into the grout holes, also starting at the bottom. This will not leak into the shield or travel far from the tunnel, but it may not fill spaces uniformly. It is therefore usually followed by regular grout. The quantity required is greatly reduced, and its tendency to travel is checked by the presence of the gravel.
Compressed Air. The compressed-air or plenum method of tunnel driving makes it possible to work with relative safety in soft mud and under bodies of water. The principle is that the inward and downward pressure of water and of soils can be counteracted by increasing the outward pressure exerted by air in the tunnel.
The rule of thumb is that each ½ pound (0.23 kg) of air pressure over atmospheric will support a 1-foot (0.305-m) height (head) of water. Actually, the pressure required is often far less, because of the stability of the soil, restriction of water passages, and other factors.
The extra pressure is built up by low-pressure compressors (converters) at the surface, and piped through a retaining bulkhead into the tunnel. Workers and materials are passed through this bulkhead through one or more locks. Air in the tunnel may leak out through the soil as fast as it is supplied, or may be exhausted from the heading through a blowline.
A lock is a passageway between two airtight doors. The outer door of the tunnel is opened to admit entering personnel or materials. It is then closed, air pressure is raised to match that in the tunnel, and the inner door opened to complete the passage.
When exiting, valves are opened to bring pressure in the lock up to that in the tunnel. The inner door is opened, the traffic moved into the lock, and the door is closed.
Air is allowed to escape from the lock until it is at atmospheric pressure. The outer door can then be opened.
This device permits maintenance of pressure in the tunnel, and limits traffic air loss to the relatively small amount in the lock at each use.
It is best practice to have at least two locks, one for people and one for materials. The human lock should be large enough for the whole crew, and must have valves by which pressure can be closely controlled so that it will drop gradually for minutes or hours while crew leaving the tunnel are in it. This process of gradual reduction, called decompression, is necessary to prevent nitrogen dissolved in the blood from being suddenly liberated to cause a painful and sometimes fatal ailment called the bends.
There may also be a small emergency lock, high in the bulkhead so as to be the last place flooded. This is left open to the high pressure, so as to be ready for immediate use. One or more cross partitions may be placed in the crown to hold air pockets in case the tunnel should be flooded.
The materials lock should be long enough to accommodate hauling units of the size used. It may be small, so that one car at a time is pushed by hand in one end and then pulled out the other; or it may be as much as 80 feet (24.4 m) long, to accommodate a train and a locomotive. Lock construction is expensive, but a liberal size speeds work greatly.
Fire danger under high air pressure is severe. The extra supply of oxygen in close contact may cause even wet wood to burn vigorously. Smoking and other fire hazards must be avoided, and there should be a liberal supply of fire extinguishers, and fire hose connected to high- pressure water.
Clay reacts most satisfactorily to compressed air, as it is so nearly impervious that it is well supported by the air and seals it in. Primary bracing may not be required before placing the permanent lining.
On first exposure to compressed air, silt acts as clay, but it then tends to dry out and crumble off at the top, and to turn to mud and flow at the bottom. The higher the tunnel, the greater the differences between top and bottom behavior.
This is because the air pressure is the same on all parts of the tunnel rim, but the head of water that tends to force water into the tunnel, or resists its being forced out of the lining soil, is much less at the top than at the bottom. A partial cure for the difficulty is to excavate the upper or arch section first under low pressure, install liner plates or other support; then increase pressure and dig the bottom. Once a full lining is installed, the unbalanced condition becomes unimportant.
In sand the air penetrates several feet at the top, and leaves the bottom wet enough that boards have to be stuffed with excelsior to stop sand runs. The best cure for this condition is to drive well points ahead of and below the face, and keep the lower sand dry until lining is placed.
Air will escape in any formation except tight clay, and will reach the surface by following porous veins, old wells, or even sewers. It is best conserved by getting the lining in immediately after the digging. Airtightness of the lining is not automatic, however. Grouting outside it (which is necessary for firmness also) and painting the inside of concrete with cement and water greatly reduce leaks.
Liner plates may be made airtight by spreading wet clay along the joints. Building paper can be used on wood lagging.
About 20 cubic feet (0.57 cu.m) of atmospheric air per minute is required for each square foot of face area, with an additional allowance for losses through the locks.
Blowouts. Sometimes the compressed air in the tunnel blows out the surface. This is particularly likely to occur in shallow tunneling in soft underwater mud. Any outward leak must be immediately plugged with any material on hand, valuable or otherwise. From the outside a blowout can be prevented or stopped by dumping enormous quantities of clay from barges.
The blowout can be disastrous in itself, hurling workers and equipment up into the water. The immediate drop in pressure allows water and mud to enter the tunnel, threatening those in it with drowning or suffocation. A job “lost” in this manner is expensive and tedious to resume, and sometimes driving can more easily be done on a different route.
In addition to this below-ground work, a variable amount of surface construction is required. The following is a list of major items of this nature:
• Access roads
Construction
Maintenance
• Power supply
Installation of lines
Construction of generating plant if necessary
• Surface buildings
Change and washroom facilities
Blacksmith shop
Machine shop
Compressor building
Powder magazine
Cap magazine
Miscellaneous buildings
• Construction camp (if needed)
• Portal excavation
• Water supply
• Sewer system
A mine is an excavation made in order to obtain (recover) material that has valuable chemical or physical characteristics. If it is an open-cut project, it may be called a pit or a quarry. Strictly speaking, a quarry is usually concerned chiefly with a material desired for its physical characteristics, as trap rock for road aggregate. However, if a quarry goes underground, it is called a mine.
The problem of mining is to get the highest possible percentage of the pay material out, at minimum expense. In some cases the best system is to confine excavation almost entirely to pay dirt, even if it requires a maze of small and irregular tunnels. In others it is more efficient to blast and remove 100 feet of overburden so as to expose only a few feet thickness of ore.
The first step in deciding upon an approach is to find how much ore or other pay material there is, exactly where it is located, the extent to which it is interrupted by other materials, and its physical and chemical characteristics. This information may be required to determine not only whether the deposit is worth mining and the method, but also the type and size of any processing plant required.
Mineral Deposits. Exploration is very complex, because of the number of factors that influence distribution of minerals. Sedimentary rocks develop in more or less horizontal layers, but may then be folded, twisted, or even turned upside down. Faults are breaks that extend across the layers, and are made by movement of whole blocks of the earth’s crust. They may be a single clean-sliding plane, or a width of hundreds of yards in which the rock is smashed up. Movement along a fault may be a fraction of an inch, so that the same formations are found on each side, or several miles, so that one section of a deposit may be at a great distance, or lost entirely. Movement may have taken place in the ancient past, or might occur from time to time during mining.
Most metallic minerals are associated with invasion of formations by molten rock from below. If fluid rock reaches the surface, it becomes the lava, ash, and other usually valueless materials associated with volcanoes. If it stays far below, it hardens gradually into granite or other coarse-grain rock, and while cooling may give off great quantities of minerals, in fluid or gaseous form, that penetrate the surrounding rock for miles. The weak or porous streaks, through which they move and in which they are deposited, are the miners’ pay veins. Parts of the main mass may become mineralized, often resulting in an extensive but low-grade ore body with rather uniform composition.
One area may undergo several successive periods of mineralization, the later ones reworking, removing, or enriching some of the earlier deposits. When the area cools sufficiently, groundwater becomes active at dissolving, transporting, and redepositing material.
The result of these factors is that underground structure is often extremely complex, and while exploration can give a general picture of what to look for, only very extensive (and expensive) diamond drilling, or the actual removal of the ore, will give the complete story. Access to formations is often obtainable only by the hardest and most costly type of digging—hard rock tunneling.
Exploration. There are a great many methods of exploring an area for valuable minerals. Until rather recently most deposits were found by surface inspection and sampling of the ground. People on foot or horseback found pay outcrops, pieces of them below or downstream, or formations associated with them. Major finds are still made in this manner, but complex techniques have become more important.
A search may start with studies of geologic maps indicating more or less completely the rocks to be found in a region. Inspection and photographs from planes may reveal promising areas, which can then be scouted on foot. Test holes can be made with almost any tool from a pickaxe to a diamond drill. Radiation-sensitive instruments such as the Geiger counter are used to locate radioactive deposits. Local changes in gravity may indicate metallic ores.
A very interesting method is seismic prospecting. A deep hole is made, usually with a diamond drill, and a heavy charge of explosives fired in it. Sensitive instruments at selected spots in the area record the time and pattern of the resulting earth waves, which indicate the nature of the ground through which they pass. The information may be used directly in locating oil and some other deposits, or in working out underground structure to indicate the location of veins whose outcrops are confusing.
Following the Ore. When mining has started, with or without benefit of thorough exploration, the digging is kept in the ore whenever it does not make too complicated a pattern. In general, large, well-financed operations are more inclined to place their haulageways and shafts with an eye toward long-term efficiency, where the small operator keeps in pay rock as much as possible, sometimes with most unfortunate effects on later operations.
Figure 9.67 shows some of the tunnels that might be included in a mine. The main route in from the portal is the haulageway. A drift is an approximately horizontal tunnel of small size, a stope is excavation of a room. A raise is a shaft worked from the bottom. Drainage tunnels, sometimes miles in length, are driven to save the cost and danger of heavy pumping inside wet highlands. The glory hole is a shaft enlarged from the top, with the muck descending by gravity to a floor from which it can be loaded by machinery. This loading area is a draw point.
FIGURE 9.67 Mine shafts and tunnels.
Any of these except the glory hole may be in either ore or nonpay rock. They are contracted to minimum dimensions (which for a haulageway may still be quite large) when in country rock, and are expanded and supplemented by side drifts when in ore. If a vein is too narrow or too low for working space, excavation may include enough other rock to give head or side room.
The “arch” of the rock (the span of roof that can be allowed between walls or pillars, with or without timbering) and the height of the veins are important factors, as they determine the yardage (called tonnage underground) that can be taken out at one stand. Larger volume and working space permits use of bigger and more efficient machinery.
Development. Development is that part of mining that prepares an ore body for removal. It is likely to include the sinking of shafts, the driving of tunnels, and the installing of chutes and transportation and drainage systems.
If development is done in the ore or other pay mineral, expenses may be partly or wholly paid by the value of the product. In some procedures, as in room and pillar mining where the pillars are left, the development may be the entire mining operation.
Shafts. Shafts are similar to those used for other tunneling. They provide for entrance of personnel and supplies, removal of ore and waste, drainage, and ventilation. They vary in size upward from a drill hole with a casing 8 inches (20.3 cm) or more in diameter that serves as a chute for concrete or fill, to openings large enough for four divisions—two for a pair of skips for bringing up ore and waste, one for an elevator for personnel, and one for pipes and ladders. Supplies may be lowered in human elevators, in ore skips, or by a separate system. There may be an additional system of ventilator shafts, that may be either dug or rotary-drilled.
Haulageways and Drifts. Tunnels that are made primarily for hauling ore or other materials from digging points to shafts or portals are called haulageways. Their size varies with that of the haulage units to be used in them. Old time human-and-wheelbarrow methods could get by with a width of 4 feet (1.22 m) and a height of 6½ feet (1.98 m). Two-way haulage with off-the-road trucks requires a width of about 30 feet (9.14 m) and a height of 20 (6.1 m). (See Fig. 9.68.)
The majority of underground mines use track haulage, although rubber-tire trucks and shuttle cars are also used. A few still use hand-push cars on 18-inch (0.46-m)-gauge tracks. Gauge is the spacing, center to center, of the rails. Locomotive haulage may call for track gauges from 24 to 42 inches (0.61 to 1.07 m). Cars may project 18 inches (0.46 m) beyond the track on each side. An additional space of at least 18 inches (0.46 m) on one side is required so as not to crush workers against the walls. There is usually a gutter or pipe for drainage water, piped high-pressure air for power and low-pressure air for ventilation, and electric wires or cable.
Haulageways that are to be used for a long time are given strong and permanent linings. Concrete without reinforcing is used because of its compressive strength, comparative simplicity of placement, and the fact that when broken, it is readily replaced or removed. However, steel lining, timber sets, roof bolting, or sometimes just shaping and scaling of a natural rock roof are all used under suitable conditions.
Drifts are tunnels made during exploration and development. They are usually smaller, shorter, or are expected to carry less material than a haulageway, but there is no clear distinction.
Remote Control of Loader/Haulers. In recent years a System for Integrated and Automated Mining (SIAM) has been developed by Noranda for their Brunswick mine. It consists of a multimedia communications backbone, a video-assisted teleoperation system, a system for automating load-haul-dump (LHD) bucket loading, and an automated guidance system for LHB machines. In other words, the remote control is operated from a control room at a safe distance from the mining face. There the operator with joysticks maneuvers an LHD into the muck at the face of the mine to scoop up a bucket full of muck, then moves the LHD to carry the muck to a dumping point. The bucket is filled by an automated system based on hydraulic pressures, cylinder extensions, and load on the front axle. With such automation the system ensures that the bucket will be essentially full every time.
FIGURE 9.68 Diesel mine hauler.
This remote control system has proven to be successful at consistently filling the bucket of each LHD at the face and displaying the load to the operator in the control room, improving productivity, and reducing costs of the mining operation. The reduction of costs is made possible by having a single operator handle several machines at a time. The system uses visual navigation and each vehicle is equipped with a series of sensors, an on-board computer, and two cameras, one at each end. The cameras transmit images through an optical tape that runs along the roof of the drift.
Extraction. Extraction is the operation of removing ore and other minerals that have already been made accessible by development work.
Minerals in extensive beds of fairly uniform thickness may be mined by a method known as “room and pillar.” Coal, limestone, and salt are frequently dug in this way.
A number of parallel corridors (tunnels or drifts) are cut into the formation, and are connected by series of parallel corridors crossing at right angles. There must be at least two of the main corridors to provide for ventilation, and there may be as many as five in a set—two haulage, two ventilation, and one spare. The corridors are the rooms, and the blocks or walls between them are the pillars.
Figure 9.69 shows a layout during the development stage, as the rooms are being cut. The rooms and pillars are the same width. This could be as little as 6 feet with a weak roof, or as great as 50 feet (15.2 m) under very favorable conditions. With such an equal-width layout, the development work would remove 75 percent of the mineral formation, if a clean cut were made to the floor and the roof.
FIGURE 9.69 Room-and-pillar excavation. (Courtesy of Ingersoll-Rand Company.)
Another method, Fig. 9.70, is to leave the pillars as continuous walls except for occasional ventilation connections between the corridors. This is often done when the corridors will be backfilled to support the roof, and the pillars then dug out.
Leaving the Pillars. The pillars may be left permanently to support the roof. This may be done because the supply of the mineral is so large that it is more efficient to waste the pillars than to do the extra work of extracting them; when roof conditions make extraction particularly dangerous; or to prevent damage by subsidence of the ground above. Worked-out mines of this type are being used for storage of records, and for underground factories.
Pillars left in this manner may be the full size left by the development, or they may be cut away (robbed) to the point where they are barely adequate to support the roof.
Limitations. Room-and-pillar work of this type is usually limited to beds whose thickness is not much greater than the safe span of the roof between pillars. Tall pillars must be thicker than short ones to support the same weight, and in thick beds with narrow corridors they will contain too high a proportion of the mineral.
A mineral body must also be fairly regular for efficient room-and-pillar work. Interruption of the pattern by barren or soft ground, or sudden changes in pitch, increases costs and decreases recovery.
The term stoping is used so differently in different mining areas that it has little specific meaning. In general, and in this discussion, it applies to any underground digging of valuable minerals that does not create a passageway. It covers most forms of ore extraction other than room-and-pillar and block caving.
A stope is the cavity in which stoping is being done or has been done. Refer to Chap. 20.
FIGURE 9.70 Four-heading cycle. (Courtesy of Ingersoll-Rand Company.)
Gravity. Almost everything is against economical work underground: restricted working space, tight blasting, need of roof and sometimes wall supports, cost of hoisting to the surface, darkness, limit on size of machinery, and drainage, pumping, and ventilation requirements.
There is one helpful force that the miner puts to use very effectively: gravity. The miner employs it to help break rock, or sometimes to do the whole breaking job, and to convey the broken material through chutes to loading points.
Chutes. Underground operations in extraction or stoping are very complex, and almost any cutaway view and brief description are likely to be incomprehensible to the surface worker. For this reason, the subject will be opened with an illustrated and oversimplified description of procedures.
If ore lying above a development tunnel is to be mined, it might theoretically be loosened by a bar or by blasting to fall directly into a mine car parked below, as in Fig. 9.71(A).
However, some of it would fall beside and between the cars, and it would be difficult to control the amount loaded. For this reason it would be better if the loosening were done somewhat higher, as in (B), and the broken ore fed down through a chute with a control gate.
But a vertical chute dumping chunks of rock out of the ceiling is dangerous, and the gate will be very difficult to work, as a great weight of rock can rest directly on it. Therefore the chute is made a sloping one coming in at the side, as in (C). And since a tunnel has two sides, two chutes can be cut into it at one point, as in (D). Several chutes or pairs of chutes may be spaced along the tunnel so that each can load a separate car of a parked train at one time. In this manner, a 12-car train could be loaded by two chute locations with six stops or by three chutes with four stops.
Gate-controlled, car-loading chutes may be as small as 2 feet square (0.19 sq.m), are unlikely to be over 5 feet (1.5 m) wide. Small sizes may be operated by hand or compressed air, larger ones by air only.
This type of chute may be very subject to jamming by oversize pieces that may be difficult and expensive to remove.
This problem is met by having a sorting and screening point between the stope and the chute, as in Fig. 9.72. A tunnel is cut above the haulage drift, into which broken ore can flow by gravity so that it will spill into the chute, or can be raked or shoveled into it. The top of the chute is protected by grizzly bars. Any piece too large to go through them is broken up, or pulled out of the way by a miner.
Chutes that will handle a considerable tonnage of ore may be lined with timber, metal, or concrete. Wear on concrete is reduced by using a stepped underslope instead of a smooth slide.
Draw Point. Jamming may also be reduced or eliminated by using a much larger chute ending in a draw point instead of a gate. The one shown in Fig. 9.73 opens into a side tunnel or room where it spills on the floor, and is loaded into cars or onto a belt by an overhead shovel or some other mucking machine. The ore is held by the floor and by its natural slope, and feeds down automatically as the toe of the pile is dug away.
Another type of draw point or free-flowing chute empties directly onto the tunnel floor, putting the ore in the working path of a drag scraper (slusher).
This machine, the two versions of which are shown in Fig. 9.74, is a bucket, usually bottomless, that is pulled to the digging area by a cable attached to its rear, and then pulled forward through its digging, transporting, and dumping cycle by the other end of the same line, attached to its front.
The line operates between a winch at or near the dumping point, and a pulley block anchored behind the digging area, several feet above the tunnel floor. (See Fig. 9.75.)
One scraper may service several draw points.
Square-Set Stoping. This method does not depend on broken ore for a working floor. The ore is cut in a succession of identical blocks, that may be between 5 × 5 × 7 and 6 × 6 × 8 feet (1.5 × 1.5 × 2.1 and 1.8 × 1.8 × 2.4 m), the largest dimension being vertical. As each block is removed, a squared timber frame is erected in the space and roofed with planks. The stability of the material will determine whether the blocks must be dug and shored one at a time, or if a number can be opened and then shored at one time.
This work may be started at the bottom of the ore body or of the section being worked, and the bottom level (or slice) is followed in sequence two or three sets behind by higher slices, so that the cross section of the front resembles the underside of stairs. The roof of each frame serves as the floor of the one above it. The line of advance may assume any one of a number of patterns, the straight wall from side to side of the ore or section, or a center advance followed by stepped backsides.
FIGURE 9.74 Drag scraper installations.
The strength of the timbering in a large stope depends largely on keeping the sets in exact alignment in all three directions. If any part starts to move, diagonals and bracing plates are installed, sections are rebuilt, or the worked-out sections may be filled with waste rock.
If the ore is mixed with country rock, a large amount of filling may be done by separating ore from waste while mucking, and moving the waste into the bottom of the stope. The convenience of doing this is a great advantage of square sets over shrinkage stoping and other caving methods that require taking everything, good or bad, to mechanical separating equipment that is usually in the mill aboveground.
FIGURE 9.75 Drag scraper anchors.
Backfilling. Square sets and other types of stope may be backfilled as a regular part of the extraction operation. This may be done to strengthen any bracing structures and support the walls and overhangs during extraction operations, to provide elevated surfaces on which miners can work, and/or to provide against long-term subsidence that might affect other parts of the mine and its equipment, or surface buildings (frequently including the mill).
The backfill may consist in part of waste separated in the stope or at the crusher, of tailings from the mill, or of granulated slag. When operated in connection with open-pit or open-stope mines, or when the proportion of waste is very high, this material may be all that is needed. Otherwise sand, gravel, or other suitable material is dug in the vicinity and hauled to the mine.
The waste fill is usually handled by a system of shafts and passages separate from that used for ore removal. Sand, boulder-free gravel, and dry tailings may be dropped through chutes made from cased drill holes. Wet tailings may be pumped down similar bores, or through pipe rigged in the main shaftway. Cement may be mixed with backfill gravel to make a very lean concrete, in order to avoid settlement and slumping into adjoining pillar extraction work.
If a large enough area of rock or ore is undermined, it will cave in. Many formations will break up in collapsing so that most of the pieces are small enough to go down ore chutes. This behavior is utilized in block caving.
The minimum width for sure caving is around 100 to 150 feet (30 to 46 m). Under various conditions block dimensions may vary from 100 × 100 (30 × 30 m) to 200 × 300 feet (60 × 91 m). Height of blocks varies from 200 feet (61 m) up to perhaps 400 (122 m), and a substantial weight of overburden may (or may not) be helpful. The block is sometimes cut off from adjoining blocks, or from separating walls by shrinkage stoping or a network of raises or drifts.
Development work prior to caving is fairly complicated. Three levels of tunnels are needed: the undercutting level where the block is undermined, the grizzly level below it for sorting and feeding into chutes, and the bottom or haulage level. See Fig. 9.76.
The levels are connected by raises. The upper set, from the grizzly to the undercut passages, are called finger raises. Their upper ends may be belled out into funnels. These terminate beside and above grizzly bar entrances to the control raises (chutes), so that oversize pieces may be broken or set aside.
When the development work has been completed for a block or a large enough part of one to start caving, the pillars separating the undercut drifts, and usually their ceilings as well, are drilled and blasted. The broken ore spreads and settles into the drifts, leaving the roof entirely unsupported, so that it breaks under its own weight and crumbles into the raises. This process may start immediately, or after a delay of several days.
The number of control raises can be greatly reduced by using drag scrapers (slushers) to pull ore away from the bottoms of the finger raises to a few large raises.
A number of problems arise in block caving. First, there is the blocking of finger raises by oversize pieces, that must be broken by “bombs” of dynamite that are pushed up the raises on poles to the obstruction, and exploded. Then there is the arching of the ore over a raise or group of raises so that it will not feed, usually a tricky and sometimes a dangerous proposition to correct.
Another difficulty is that uneven breakage and settlement may put great pressure on certain sections, crushing even heavily reinforced control and haulage tunnels. If the blocks are separated by pillars (walls), and the barren capping rock is stronger than the ore, settling in the blocks may result in these walls supporting large areas of overburden, with resultant heavy pressure on passages under them. Another possibility is that a large mass of rock in one of the blocks may be more resistant to breakage than the ore, so that it rests more and more heavily on the ground beneath it as the ore crumbles and is drawn away. One block-caving operation lost a haulageway for 18 months as a result of such a mass crushing it repeatedly.
In spite of these and other complications, block caving is the most economical method of mining large, fairly uniform ore bodies that will break into small pieces under their own weight, and that are too far underground for open-pit recovery.
Surface Subsidence. On the planning side, any mining method that allows the roof to cave as the ore is taken must allow for the effects on the surface. A thin, even bed such as a 6-foot coal seam may let the surface down evenly, with comparatively little damage to structures. But any substantial and irregular removal of underground material will result in subsidence pits. These may be hundreds of feet deep and thousands of feet wide, and are likely to destroy any road or structure within their reach.
The subsidence of a caving block may be nearly vertical at first, but the cleavage in the overburden is likely to fan out widely. The ultimate effect, in a wide variety of rock formations, is crumbling and sinking inside a slope line of 40° to 45° outwardly from the bottom of the block or caving operation, with additional slip faults that might extend to a 38° slope. See Fig. 9.77.
All too often ore treatment mills, shaft head structures, and towns are involved in such subsidence, so that very expensive relocation is required.
One of the principal safety problems of underground work is the danger of rockfalls and rockbursts. Knowledge of where such incidents are likely to occur usually makes it possible to install additional supports, or if that is not possible, to remove personnel and equipment from the danger area.
The cracking and popping of rock, often heard by personnel working underground, has long been recognized as a warning of moving, unstable ground. It has now been discovered that sounds of the same type, so slight as to be heard only by very sensitive instruments, are made by ground that is under even moderate stress. Increasing frequency and intensity of the sounds, called microseisms, indicates an increasingly unstable condition that is likely to result in collapse.
FIGURE 9.77 Surface subsidence.
Microseisms occur within the audio frequency range and, when amplified, sound like “clicks” akin to the creaking of wooden stairways, floors, and diving boards. Almost any hard material, such as salt, rock, shales, sandstone, quartz, brick, glass, wood, steel, concrete, sand, and sugar, will produce microseisms. Microseisms by their clicking nature can be easily distinguished from sounds caused by foot shuffling, drill rigs, trucks, mucking cars, talking, and other means. Microseismic apparatus can never be used when vibrations are so great as to overwhelm the equipment.
The Seismitron (Fig. 9.78) is an instrument based on the microseismic principle and has been in use since 1951. Its primary purpose is for the prediction of rockfalls in tunnels and mines as well as for the determination of stability of earth dams and slopes at the sides of mountains and highways. The apparatus is made by the Walter Nold Company of Natick, Massachusetts, and is certified (permissible) for use in coal mines by the U.S. Bureau of Mines and by the Department of Mines and Technical Surveys for Canada.
Receiving sensors (geophones) are cylinders 1¼ inches (3.18 cm) in diameter and 9 inches (22.9 cm) long. They contain synthetically grown crystals which, when stressed, generate low-level electric currents. A portable battery-powered amplifier increases the level to a condition that enables the microseisms to be heard by means of simple earphones.
The number of microseisms per minute which occur when the rock is at the failure point, or about to burst, is determined by monitoring the heading (of a tunnel, for instance) immediately after a blast. The microseismic rate of decay under such a state is the same as the rate of increase prior to a rockfall. This is the failure point.
Six-foot (1.8-m) deep holes are drilled and spaced every 100 feet (30 m) apart in the suspected area. Each of these stations is monitored each week for microseismic activity. This is accomplished by placing the geophone in the hole and sealing the opening with waste. The microseisms heard are manually counted for a 15-minute period and the rate per minute versus dated time plotted on graph paper at a later time.
A history is thus built up for each station. A low, little-changing rate of microseismic activity is indicative of stable conditions. A continually increasing rate indicates suspect conditions, especially when 120 to 180 microseisms occur per minute. A doubling of the rate within any 24-hour period indicates imminent collapse.
An excellent prediction of collapse, sometimes as much as 45 days ahead of time under ideal conditions, can be made by extrapolation of the plots on the graph to the rock failure point. The curvature becomes hyperbolic as failure approaches.
If the active rock, which is sometimes hidden by gunnite, should become stabilized either naturally or by roof bolting or timbering, the decaying microseismic rate would immediately indicate the tendency toward safe conditions.
The instrument in this manner becomes a very useful tool for judging the necessity for, or the effectiveness of, expensive roof bolting. The Seismitron is used in earth drilling to determine depth of bedrock. It is also used to determine leaking running water in earth dams, evidence of life in blocked tunnels and earthquake rubble, stability of overhanging cliffs, large ice masses, concrete structures during quiet periods of construction, underground foundations while underpinning is being replaced, communication through rock or pipe, robbing pillars of mineral-bearing ore, and so forth.
With the Seismitron, timbers about to break sound just as one would expect wood to sound under these conditions. Rock about to fall sounds like rock about to fall. Earth about to slide sounds just like earth about to slide. In other words, just like the subject of physics; natural and normal.
