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Rock Blasting in Underground Mining
Chapter 19
Rock Blasting in Underground Mining Chapter Outline 19.1 19.2
19.3
19.4
19.5
19.6 19.7 19.8
Advantages and Disadvantages of Sublevel Caving Drilling and Charging Plan 19.2.1 Burden and Spacing 19.2.2 Length of Blastholes 19.2.3 Uncharged Length 19.2.4 Precision of Rock Drilling Inclination of Rings 19.3.1 Loading and Ore Flow 19.3.2 Brow Protection Single- or Multiple-Ring Blasting 19.4.1 Single-Ring Blasting 19.4.2 Double-Ring Blasting Delay Time and Initiation Sequence 19.5.1 Delay Time 19.5.2 Initiation Sequence 19.5.3 Initiation and Delay Time in Special Cases Open Cut and Drifting Stemming, Air Deck, and Detonator Placement Back Break and Brow Damage 19.8.1 Back Break 19.8.2 Brow Damage
373 374 374 375 377 377 377 377 379 380 380 381 382 382 382 382 383 383 384 384 385
19.9 Misfires 19.10 Fragmentation, Ore Recovery, and Mining Profit 19.11 Safety, Environment, and Vibration Control 19.11.1 Safety 19.11.2 Environment 19.11.3 Ground Vibrations 19.12 Sublevel-Caving Blasting in the Future 19.13 Concluding Remarks 19.13.1 Characteristics of Sublevel Caving Blasting 19.13.2 Subdrilling, Deviation, and Inclination 19.13.3 Single-Ring Blasting and Double-Ring Blasting 19.13.4 Delay Time 19.13.5 Brow Damage 19.13.6 Misfire 19.13.7 Other Issues Related to Sublevel Caving Blasting 19.13.8 Underground Blasting in the Future 19.14 Exercises References
387 387 388 388 388 389 389 390 390 390 391 391 391 391 391 391 391 392
There are clear differences between production blasting in underground mines and that in open pit mines. First, in many underground mines a production blast consists of a fan-shaped blastholes in each row. Such a row is called a fan or a ring, which appears in different mining methods such as sublevel caving, block caving, and open stopping. The blastholes in a ring are not parallel to one another. Second, in most underground mines production blastholes are often drilled upward instead of downward as in open pit mines. The upward blastholes make it difficult to apply stemming. Because the sublevel caving method is widely used in various types of underground mines, we will focus on the blasting in sublevel caving in this chapter. However, the principles in sublevel caving blasting are applicable to other underground mining methods. Rock blasting in sublevel caving deals with drilling plan, charging plan, quantity of rings in each blast, back break, brow damage, and misfires. These will be described in detail in this chapter. In addition, we will outline delay time, explosive initiation, open cut, drifting, stemming, detonator placement, relation between fragmentation and ore recovery, and effects of blasting on safety and the environment since they are described in detail in other chapters. Last, blasting in sublevel caving in the future will be discussed.
19.1 ADVANTAGES AND DISADVANTAGES OF SUBLEVEL CAVING Sublevel caving is a widely used mining method in metallic mines all over the world. The modern sublevel caving was possibly developed in the iron mines of Sweden [1]. This method has many advantages with regard to safety and mechanization [2]. Because all of mining operations are carried out only in drifts, the safety in sublevel caving is relatively good, compared with other mining methods such as cut and fill, room and pillar, and so on. In addition, since big machines can be used in sublevel caving and mechanization and automation are available, mining production can be performed on a very Rock Fracture and Blasting: Theory and Applications Copyright © 2016 Elsevier Inc. All rights reserved.
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large scale. For example, automation or remote control in production drilling, underground transportation, extraction, and the like has been realized for decades in a number of large mines such as Kiruna and Malmberget mines. However, sublevel caving has some disadvantages. High ore loss and high dilution are two of them. It is not unusual for a sublevel caving mine to have ore loss up to 20% in the mining process. In brief, the major advantages and disadvantages of sublevel caving can be concluded as follows. The major advantages are: l
Good safety; High mechanization and automation; l Large-scale production and high productivity. l
The major disadvantages are: l
High ore loss; High dilution.
l
Therefore, how to eliminate these disadvantages is a challenging task for the sublevel caving method. Fortunately, production tests have shown that ore loss can be markedly reduced by improving blasting. This will be presented in the chapter: Effect of Blasting on Engineering Economy.
19.2 DRILLING AND CHARGING PLAN 19.2.1 Burden and Spacing 19.2.1.1 Burden How to determine burden in rock blasting has been described in the chapter: Burden and Spacing. Burden is dependent on more factors in underground mining than in open pit mining. For example, burden is related to mining scale, which depends on the sizes of equipment such as loading and drilling machines; burden influences ore recovery in sublevel caving, too. For instance, if burden is too large, it is difficult for a loading machine to reach the ore in the farthest muckpile. This makes part of the ore in the muckpile remained. In this case, if the ore body is narrow or only one drift is on each level, the remained ore (or at least part of it) will become a permanent loss. In addition, burden affects vibrations. In brief, the following factors are related to burden: l
l l l l
Production scale; Sizes of drilling, charging, and loading machines; Fragmentation; Ore recovery; Ground vibrations.
In blast design of sublevel caving, it is suggested that the burden should be determined on the basis of the analysis in the chapter: Burden and Spacing combined with empirical parameters. In present sublevel caving, the largest borehole is about 115 mm in diameter, and corresponding burden is often 3 m or 3.5 m in some mines. These parameters may be taken as a reference but not necessarily as the best choice since no study indicates they are optimum.
19.2.1.2 Spacing Drilling and blast plans in sublevel caving are different from those in open pit blasts. As shown in Fig. 19.1, a ring in sublevel caving looks like a fan. Note that the spacing in the sublevel ring is not a constant but the spacing in open pit blasting always is. From Fig. 19.1 we see that in the upper part of the ring the spacing is much larger than it is in the lower part; that is, the spacing for two neighboring holes varies from their collars to bottoms (eg, the collar of borehole 4 is at C). In addition, the boreholes all have different lengths and different inclinations. If only one kind of explosive is well distributed in a borehole, that is, the explosive and its density do not vary along the borehole, the varying spacing has a negative impact on fragmentation since the upper part of the ring always has a lower specific charge. This makes the ore in the upper part poorly fragmented, compared with the ore in the lower part. Consequently, in sublevel caving it is often the case that as soon as extraction starts after a ring is blasted, the sizes of fragments increase with increasing ore extraction, meaning that boulders usually come from the upper part of the ring. To reduce boulders or improve fragmentation, the spacing in the upper part must be reduced. In this way, the uncharged length in the lower part of every next hole must be increased in order to avoid misfires. Two other alternatives are that (1) in the upper part of ring a stronger explosive with higher velocity of
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FIGURE 19.1 Production ring in sublevel caving. Rf is the radius of the fragmentation region of each blasthole; Sch is the distance from the end of the explosive charge in one hole to another that is nearest. There should be Sch ≥ Rf if delay time is not very short. The uncharged part of hole 6 is too short. The correct position should be at the other end of the short line at I.
detonation (VOD) is charged, and (2) one ring is divided into two parts—upper part and lower part—in blasting and a very short delay time is applied in the upper part to make full use of stress wave superposition [3]. Although the first alternative is difficult to implement in production today, it will probably be more possible in the future. An introduction to the second alternative will be made in the chapter: Reduction of Ground Vibrations. Up to now boulders have been a big problem in underground mining. On one hand, the current level of specific charge, which is often 1 kg/m3 more or less, is too small in underground mining, indicating that the specific charge may have a great space to increase from the present level. On the other hand, the blasting methods in current production blasting have much to improve. For example, detonator placement, stemming, delay time, and the like all need a scientific design.
19.2.2 Length of Blastholes As soon as the sublevel height is determined, the size of a ring and the length of blastholes in a ring can be decided. We discuss two cases: one for a large ore body and the other for a narrow ore body.
19.2.2.1 Large Ore Body In a large ore body, there are two types of rings: (1) the rings under the hanging wall that is composed of waste rock mass, as shown in Fig. 19.2a, and (2) ordinary rings as shown in Fig. 19.2c. In ordinary rings, there must be a zone called undrilling or buffer left in the top of the ring, as shown in Fig. 19.2c. In other words, a length indicated by “Undrilling” is left without drilling. If the boreholes in a ring reach its upper boundary, after production in the upper level is finished, some fragments
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FIGURE 19.2 Subdrilling and undrilling in wide ore body. (a) Rings under hanging wall; (b) large ore body with hanging wall; (c) ordinary rings.
may fall into and block the boreholes. This will make it impossible to charge explosive in the upper part of the boreholes. As a result, fragmentation will be worse. Therefore, we need the undrilled part. In sublevel caving 1 m or so can be tried for the length of undrilling. If it is too long, the fragmentation will be worse. If the ring is under a solid hanging wall, as shown in Fig. 19.2a and b, the subdrilling is necessary, like in open pit drilling. Otherwise, without subdrilling, that is, the bottoms of the boreholes are at the upper boundary of the ring after blasting, the ore mass surrounding the bottoms will not be completely broken since confinement is strong there. This is similar to blasting in open pit mines. As a consequence, that part of ore mass will be left as a permanent loss. In order to avoid such an ore loss, a subdrilling is necessary. It is suggested that 0.5–1.0 m can be taken as a reference for the size of subdrilling in sublevel caving.
19.2.2.2 Narrow Ore Body In a narrow ore body, there is often one drift on each production level, as shown in Fig. 19.3. Several boreholes close to a hanging wall must have subdrilling so as to avoid ore loss near the boundary between ore body and hanging wall. In contrast, without subdrilling, the ore mass close to the boundary will not be well broken due to high confinement surrounding
FIGURE 19.3 Sublevel ring in a narrow ore body.
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the bottoms of boreholes. Finally, this ore mass will be left as a permanent loss. Such a permanent ore loss may be considerable if a mine has many narrow ore bodies and its production is large.
19.2.3 Uncharged Length In a ring blast, the length of the uncharged part of a borehole must be carefully determined due to the varying spacing. Such a length is mainly dependent on the fragmentation radius Rf of each hole; see Fig. 19.1. There are two different situations: (1) as delay time is longer than the duration of fragmentation propagation between the two holes, and (2) as delay time is shorter than the duration. These are to be discussed in the following.
19.2.3.1 As Delay Time Is Longer Than Fragmentation Time Under this condition, as shown in Fig. 19.1, if the distance Sch is larger than the fragmentation radius Rf of the first borehole, detonation of the first-initiated hole will not destroy the second-initiated one. In contrast, if the distance Sch is smaller than Rf, the detonation of the first-initiated hole will destroy the blasthole to be initiated immediately following the first, and consequently the explosive in the latter (second) will not detonate. As a conclusion, when delay time is longer than the fragmentation time between two neighboring holes, the distance Sch must be equal to or larger than Rf. If so, the detonation from one borehole will not influence or destroy its neighboring holes. According to the experience of LKAB, the fragmentation radius Rf is approximately 1.5 m when the borehole diameter is 115 mm and emulsion explosive is used. For other mines, it is better to determine Rf by carrying out their own tests.
19.2.3.2 As Delay Time Is Shorter Than Fragmentation Time In this case, the distance Sch can be shorter than the fragmentation radius Rf since the second-initiated hole will start to detonate before fragmentation from the first-initiated hole reaches the second hole. This will happen when electronic detonators with a short delay time are used. In this case, efficient stress wave superposition from neighboring holes may be realized. This case is an attractive area for production blasting in both underground and surface blasting in the future. In order to achieve a successful blast result, it is necessary to measure detonation wave or stress waves in the rock. Based on this, a correct delay time can be determined. Otherwise, if the determination of delay time is not based on detonation wave measurement and stress wave analysis, the delay time that is chosen may not be correct. As a consequence, the blast result will not be very good.
19.2.4 Precision of Rock Drilling The deviation of rock drilling in sublevel caving can be divided into three types, as shown in Fig. 19.4. (1) The borehole curves in two directions from its original central line, as shown in Fig. 19.4a and b. (2) The angle of the borehole changes but the collar position does not vary and the borehole does not curve, as shown in Fig. 19.4c. (3) The collar position is changed and the borehole is curved in only one direction, as shown in Fig. 19.4d. Any type of deviation shown in Fig.19.4 makes rock fragmentation worse. For example, two extremes of deviation in Fig. 19.4a are (1) the bottom distance between two holes 4 and 5 is much larger than that planned, which may produce boulders there; (2) two holes 5 and 6 are closer to each other, meaning that either sympathetic detonation occurs if they are too close, or one is destroyed when the other is detonated. The latter will also happen if the deviation is like in Fig. 19.4c. This will occur when pyrotechnic detonators are used since their delay time between two adjacent holes is long. The similar misfire will happen if the deviation in Fig. 19.4d appears. Therefore, precise drilling is important in sublevel caving as well as in other similar blasting.
19.3 INCLINATION OF RINGS There are usually two types of ring inclinations: (1) inclined rings as shown in Fig. 19.5 and (2) vertical rings as shown in Fig. 19.6. We will see the differences between the two cases. Assuming the sublevel height, drift sizes, burden, borehole diameters, length of stemming (or uncharged length), explosives, and other parameters are all identical for the two cases, we will compare the two different inclinations in the following.
19.3.1 Loading and Ore Flow If the widths of the muckpiles in both cases are equal, as represented by Wf in Figs. 19.5 and 19.6, the maximum distance available for loading along the drift axis will be the same for the two cases. In addition, the contact areas between waste
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FIGURE 19.4 Different types of deviation. (a) Deviation of borehole 5 occurs in the plane of the paper, while the deviation of other holes happen in the plane perpendicular to the paper plane. (b) Cross section of (a). (c) The collar position of hole 3 is not changed but its angle is altered. (d) The collar position of hole 5 is changed and the hole is deviated.
FIGURE 19.5 Inclined ring.
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FIGURE 19.6 Vertical ring.
rock and fragmented ore on the top of the rings are equal in the two cases, meaning that ore dilution caused by the mixture of waste rock and ore from the tops of the rings will be the same. If the burden in both cases is constant in the horizontal direction, the inclined ring and the vertical one will produce the same amount of ore, as shown in Figs. 19.5 and 19.6. For ore flow, it is difficult to find a marked difference between the two cases.
19.3.2 Brow Protection For brow protection, the inclined ring is better than the vertical ring, according to the following analysis. Under otherwise identical conditions, we assume that detonation direction is up–down and the velocity of detonation is greater than the velocity of the P-wave in the ore or rock mass. In this case, when the last part of the explosive charge above stemming is exploded, the front of the P-wave from blasting will first reach both the front surface and the roof of the drift, as shown in Fig. 19.5. After the P-wave arrives at the surfaces, two tensile waves will be caused by the reflection of the P-wave. Tensile rock fracture will start within the region surrounded by the front of the tensile wave (notice that there is also a compressive P-wave there, but the resultant stresses will mostly be tensile). Therefore, we can further assume that the size of brow damage is proportional to the size of the tensile wave region. In other words, the larger the tensile wave region is, the larger is the region of brow damage. As shown in Figs. 19.5 and 19.6, after the same time when detonation at the point O occurs, the front of the P-wave arrives at the circle indicated by the front of the compressive P-wave. Let the radius of this circle be R. In the inclined ring we assume that the brow damage Bebi is equal to the distance from G to H (in reality Bebi must be smaller than GH), and the length Lst of stemming is the distance between O and G. Since the angle of the inclined ring is bb, we can find: LFG = Lst cos β b LOF = Lst sin β b where LFG is the length of FG and LOF is the length of OF. Then we may get that the brow damage Bebi in the inclined ring is equal to
Bebi = R 2 − ( Lst sin β b )2 − Lst cos β b
(19.1)
In the vertical ring we can easily see that the corresponding brow damage Bebv is equal to
Bebv = R 2 − L2st
(19.2)
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By comparing Eqs. 19.1 and 19.2, we can find that as long as bb < 90 degrees, we have Bebv > Bebi
(19.3)
This formula indicates that the vertical ring produces larger brow damage than the inclined ring.
19.4 SINGLE- OR MULTIPLE-RING BLASTING In sublevel caving, one blast often contains one ring, but in some mines one blast consists of two rings. What are the differences between the two methods, and which one is better? This will be discussed in the following.
19.4.1 Single-Ring Blasting Before a production ring is blasted, the front face of the ring is always against waste rock (often boulders) with a mixture of the ore fragments remained from previous blasts. In a normal condition, the waste rock as well as the ore fragments remained can freely flow, meaning that the front face is a partial free surface. Therefore, a compressive stress wave can partly go through the face into the caved waste rock. The major advantage of the single-ring blast method is that the blasted ore can be extracted as much as possible. To make the analysis simple, assume that the farthest position which the bucket of a loading machine can reach is F in Fig. 19.7. Notice that OF is a vertical line. At the same time, assume that the rest angle of flowing ore is brest. Thus, the ore within the area E–G–F–H–E cannot be loaded out, and it will be remained. In a narrow ore body, that ore may partly become a permanent ore loss. As shown in Fig. 19.7, when drift height is Hdrift(= OF) and the width of ore flow is Wflow(= HO), we have H drift = tan β rest Wflow This equation can be used to estimate the amount of the ore remained. When Hdrift = 5.5 m and brest = 60 degrees, we get Wflow = 3.17 m. As shown in Fig. 19.8b, some ore close to the foot wall may be left if the angle bfoot of the foot wall is smaller than the angle brest of the flowing ore. If both bfoot and brest are known, the amount of ore left can be calculated. This part of the ore is almost impossible to recover from a lower production level (level 2 in Fig. 19.8b). In order to reduce such ore loss, the drilling plan should be made carefully, for example, to let the distance between the last hole (ie, first hole from the right side of
FIGURE 19.7 Ore extraction in single-ring blasting. Ring Ri has been blasted but ore extraction has not started yet. Ring Rj is to be blasted next time, and its front face is at AO, which is a partial free surface either during or after extraction.
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FIGURE 19.8 Ore loss. (a) Loss in large and wide ore body; (b) loss in small and narrow ore body.
Fig. 19.8b) and the foot wall be as small as possible. In addition, in the loading operation, the loading machine may load only the ore near the foot wall at the drift in the beginning of loading. This can make the ore close to the foot wall flow down easily. All in all, for a narrow ore body as shown in Fig. 19.8b, it is better to use the single-ring blasting method in order to reduce permanent ore loss.
19.4.2 Double-Ring Blasting As for two rings that are blasted together in one blast, we discuss two cases: one for a narrow ore body and the other for a large ore body. In a narrow ore body as shown in Fig. 19.8b where there is only one drift on each production level, the remained ore shown by the area E–G–F–H–E in Fig. 19.9 is much more than that in Fig. 19.7 where one ring is blasted. As
FIGURE 19.9 Ore extraction in double-ring blasting. Two rings R1 and R2 are blasted together in one blast.
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discussed previously, if a narrow ore body is not inclined very steeply, the ore remained in the narrow ore body cannot be wholly recovered in the lower level. Therefore, the double-ring blasting method is not a good choice for narrow ore bodies. In a large ore body as indicated in Fig. 19.8a, the situation is different from that in a narrow ore body. The ore left in most rings, excluding the rings near the foot wall, on a production level (eg, Level 1 in Fig. 19.8a) can be wholly extracted in the lower levels, theoretically. Thus, the ore loss from double-ring blasting will not be larger than that from single-ring blasting. In addition to ore loss, we will talk about fragmentation and dilution in double-ring blasting. On fragmentation, for double-ring blasting, the throw of fragments from the second ring is largely hindered by the first ring, making a collision happen between the fragments from the first and the second rings and resulting in secondary fragmentation of these fragments. As a result, the fragmentation in the second ring is often better than that in the first ring or the single ring. This is consistent with the field observations and production tests in open pit blasting introduced in the chapter: Rock Blasting in Open Pit Mining. Since fragmentation in a double-ring blast is generally better than that in a single-ring blast, the dilution in the doublering blast may be reduced through controlling ore extraction. For example, the low ore loss and low dilution method tested in China [4] is to leave a certain amount of ore in each production level so as to always form a layer of ore over the ongoing production level. By doing this, the interface area between the ore and the waste rock can be greatly reduced, and finally the dilution will decrease. This method is considered to be efficient in large ore bodies. Therefore, the double-ring blasting method together with an effective control of ore extraction is worthy of testing in sublevel caving with large ore bodies.
19.5 DELAY TIME AND INITIATION SEQUENCE 19.5.1 Delay Time The principles for determining delay time have been described in the chapter: Delay Times, and they are valid for sublevel caving blasting. As described in that chapter, in order to obtain optimum delay time we have to know the stress waves in the rock. Up to now, the applications of electronic detonators have not shown a much better result such as improved fragmentation in some mines. One of the major reasons is that the delay time used is out of efficient wave superposition. As soon as the stress wave (or detonation wave, which can be taken as a reference) is known, the principles in the chapter: Delay Times can be used to determine a correct delay time.
19.5.2 Initiation Sequence In normal conditions of sublevel caving, the initiation in a ring blast starts at the middle blasthole such as No. 4 in Fig. 19.10 since it is thought that the confining pressure in front of the middle hole is smaller than that of others. However, in some cases, for example, as ground vibrations caused by blasting must be controlled, the initiation sequence needs to be changed, rather than just starting at the middle hole. That is to say, if needed, the initiation may start from any hole except for the first one (eg, No. 1) and the last one (eg, No. 7). This has been proved by the vibration control method used in the Malmberget mine, where the initiation starts from the blasthole nearest to either the first or the last hole [5]. Optimistically, the production blasts indicate that the new initiation sequence (starting from any holes except for the middle, first, and last ones) does not adversely affect rock fragmentation.
19.5.3 Initiation and Delay Time in Special Cases As mentioned previously, deviation is a common phenomenon in rock drilling. It is quite often that the deviation occurs in one or more boreholes in a ring, as shown in Fig. 19.10 where borehole 5 has a deviation. The deviation of borehole 5 increases the distance between 4 and 5 at their bottoms and decreases that between 5 and 6. In this case, if the initiation starts from borehole 4 and the delay time between two neighboring holes is normal (ie, not instantaneous or very short), the consequence will be (1) that boulders are produced in the upper ring between boreholes 4 and 5, and (2) that when borehole 5 is blasted, borehole 6 may be destroyed at the same time. This happens since the Sch between holes 5 and 6 is smaller than Rf. In order to avoid this, the initiation and delay time in the ring can be adjusted. For example, when the initiation sequence is not changed, the delay time between boreholes 5 and 6 should be very short or even simultaneous so that both holes are safely blasted.
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FIGURE 19.10 Initiation sequence in a sublevel ring. When hole 5 is fired prior to 6, the EF part of hole 6 will be destroyed since Sch < Rf.
19.6 OPEN CUT AND DRIFTING Open cut and drifting have been described in the chapter: Rock Blasting in Open Cut and Tunneling. How to make a plan for open cut and drifting in sublevel caving may follow the principles there.
19.7 STEMMING, AIR DECK, AND DETONATOR PLACEMENT Stemming, air deck, and detonator placement have been described in the chapters: Stemming and Change Length; Air Deck and Smooth Blasting; Primer Placement. Here we just want to address some points. First, stemming must be applied in underground mining, even though an ideal stemming is difficult to implement in underground mines. As described previously, both material and sizes of stemming affect blast results. Therefore, it is important in rock blasting to choose a proper material and correct sizes of stemming. Second, air deck and other charge techniques may be employed in underground mines if necessary. Third, if boreholes are short, one detonator or primer is enough to achieve a good result in blasting. In this case, the best detonator position is at the middle of the charged borehole (not the middle of the total length of the borehole) with a correct stemming. If boreholes are long and multiple-detonator positions are available, two detonators positioned in each hole may be well applied. If the two detonators at different places in a hole are initiated simultaneously, shock collision will happen and it will give rise to the final shock pressure being more than the sum of the two shocks. This will to a great extent strengthen rock fragmentation. This double-primer method has been tested in the Malmberget mine with a result of less
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brow damage and higher recovery [3]. Moreover, no damage to the structures nearby has been found due to this method. However, as the double-primer method is used in sublevel caving, ore flow may change markedly, depending on the delay time. Accordingly, a prestudy or a pretest is necessary so that a correct blast design is made finally.
19.8 BACK BREAK AND BROW DAMAGE 19.8.1 Back Break Back break and brow damage are common in sublevel caving, but they are sometimes the same thing. Therefore, a definition is needed for both. As shown in Fig. 19.11, part of the burden in ring R2 is broken off by the blast of R1. The average thickness of this part of the burden along the whole ring face is called back break, indicated by Bb. However, brow damage, indicated by Beb, means that part of the burden at the roof of drift is broken off. Brow damage is often larger than back break in sublevel caving. It is difficult to measure back break in the field. When a blasthole in a sublevel ring such as R1 in Fig. 19.11 is blasted, radial cracks will be produced surrounding the blasthole. Then when the compressive stress wave propagates to the front surface of the ring (ie, a partial free surface), it is reflected into a tensile wave. This tensile wave will travel toward the rings behind R1, such as R2 and R3, since the rock surrounding the blasthole needs enough time to be completely fractured. After the tensile wave passes R1, the rock behind it will start to fracture if the tensile stress is larger than the dynamic tensile strength of the rock. In this way, the rock behind R1 is broken; that is, a back break is formed. A back break is a typical spalling caused by a tensile stress wave as described in the chapter: Stress Waves. Remember that the compressive stress wave also contributes the back break, since it causes radial cracks surrounding the boreholes. To reduce the back break, the amplitude of the tensile stress waves must be reduced. This can be achieved by reducing the size of the blasthole. When doing this, one needs to ensure that fragmentation will not be worse. The factors influencing back break are sizes of blasthole, specific charge, VOD, primer placement, initiation plan, blast plan (single-ring or multiple-ring blasting), rock properties, and geological conditions. Other measures for reducing back break can be developed on the basis of the aforementioned factors. For example, by using a double-primer placement, brow damage has been reduced [3]. Probably, back break may be reduced by the same method. Back break in open pit blasting is quite similar to that in underground blasting, as measured by Olsson et al. [6].
FIGURE 19.11 Back break and brow damage.
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Since direct measurement of back break for whole blastholes in underground mining is difficult, we usually see the back break from the roof of drifts, as shown in the next section.
19.8.2 Brow Damage 19.8.2.1 Consequences Caused by Brow Damage Brow damage is also called brow break or eyebrow break in sublevel caving. Serious brow damage decreases ore recovery and increases dilution, and also brings about worse safety for the people working in the field. For example, miners must go to the caving area and stand under the broken brow. Brow damage is caused mainly by two parts of tensile stress waves: one from the front face of a blasting ring, and the other from the roof of the drift. We will make a detailed analysis in the following.
19.8.2.2 Factors Influencing Brow Damage Main factors affecting brow damage are detonator position, quantity of primers in a blasthole, specific charge, VOD, initiation plan, contact condition of the front face of the rings, and geological conditions of rock/ore.
19.8.2.3 Measures for Reducing Brow Damage The measures for reducing brow damage are correct detonator position, multiple-primer placement, decoupled charge, higher VOD, and stopping tensile stress waves such as by presplit blasting. In the following we will briefly introduce these measures. First, detonator position, as analyzed in the chapter: Primer Placement, largely influences brow damage in sublevel caving. Fig. 19.12 shows that when a detonator is placed at the middle of a charge, the brow damage is greatly reduced, compared with the damage when a detonator is placed near the collar. Second, on the double-primer method discussed in chapter: Primer Placement, as shown in Fig. 15.6 , when the explosive in the double-primer hole is completely detonated, the stress waves with all the energy are just distributed within the rock to be broken, while no wave comes to the brow area. From this point of view, the double-primer method is as good as placement of a single middle primer regarding brow protection. One more advantage of the double-primer method, compared with the single middle primer method, is that a shock collision happens, which results in fast rock fracture in the area between the two primers, as shown in Fig. 15.6. This fast fracture will mean there will be less of a tensile wave traveling to the brow area. In consequence, brow damage is reduced, as shown in Fig. 15.12, which indicates the result from electronic detonators [3]. As a matter of fact, brow damage can also be reduced by the double-primer method with Nonel detonators, as shown in Fig. 19.13.
FIGURE 19.12 Detonator position and brow damage. (a) When detonators were close to the collars, there was large brow damage and the burden of next ring (in charging process) was seriously broken off. The brow damage even extended to the ring behind the ring to be blasted next time; see the black area to the right of the dashed line. (b) When detonators were at the middle of charged holes, there was little brow damage and most of the burden of the next ring to be blasted was remained with no damage.
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FIGURE 19.13 Photographs of brow damage taken from two neighboring drifts in an ore body at Malmberget mine. An ordinary charge was in (a) with one primer at 15 m away from the collar, and a double-primer charge was in (b). The detonators were Nonel.
Third, we will show how to reduce brow damage by a decoupled charge (ie, using a cartridge charge). This method has been successfully tested in the Malmberget mine [7]. The principle of the method is simple: to reduce the tensile stress at and near the brow. In order to implement this, some holes, like No. 3 and No. 5, are charged with cartridge explosive (ie, a decoupled charge), as shown in Fig. 19.14. In this way, the amplitude of the tensile stress waves in the brow area decreases, and finally the brow damage can be reduced. The tests in the Malmberget mine indicated that the brow damage had been reduced from 3.0 to 2.3 m by using a cartridge charge [7]. Fourth, if the VOD of the explosive is very high, for example, it is much higher than the sonic velocity of the rock, the stress distribution surrounding the borehole will soon become more even. This may partly reduce the tensile stress wave coming to the brow area. Fifth, presplit blasting can be made between two rings in order to reduce brow damage. However, notice that this may markedly increase the cost and working hours for blast operation. In addition to these five measures, there are possibly other methods for reducing brow damage. This needs further study. An interesting result is that the ore recovery from both the single middle primer rings and the double-primer rings is higher than that from the ordinary rings according to the tests in the Malmberget mine [3,7].
FIGURE 19.14 Cartridge charge in the lower parts of boreholes.
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19.9 MISFIRES Malfunction or misfire is quite common in underground blasting. According to field measurements in open stopping operations, the misfires were up to 30% [8]; in drifting, the misfires were about 23% [9]; in sublevel caving production blasting, the misfires were from 10 to 18% in the Malmberget mine. The lower percentage was from the middle primer rings and the higher one from the lowest primer rings. Anyway, misfire is still a serious problem in underground mining. There are a number of factors causing malfunction or misfire: 1. Pressure transients. It is widely suspected that pressure transients, generated by detonating explosives and transmitted through the surrounding rock or stemming material, can cause significant malfunctions in both initiation systems and explosives [10]. Mencacci and Farnfield showed a picture of two detonators that were severely damaged by pressure transient in a multidecked blasthole [10]. This may happen in the production blastholes that are precharged with explosives and detonators. In other cases where the P-wave speed of rock is greater than the VOD of the explosive, malfunction of the explosive may occur since the explosive and detonator in front of the shock waves may be highly compressed by the stress waves from the rock before the shock waves from detonation arrive. 2. Too-short distance between two neighboring blastholes; that is, the distance between the two neighboring holes is smaller than the maximum fragmentation radius of the earlier detonated hole, and the delay time between the two holes is longer than the critical time during which rock fragmentation from the first-initiated hole can reach the lately initiated hole. This case often occurs when detonators are placed near the collars of blastholes according to the field measurements in sublevel caving [5]. 3. Sympathetic initiation of neighboring holes due to fractures or cavities within rock mass between the holes. In this case the detonation waves can directly travel from the detonating hole to its neighbor and initiate the explosive there. 4. Too-long delay time between two neighboring holes. A very long delay time makes cracks from a detonating hole propagate longer. If a number of such cracks reach the neighboring hole, then the latter will be destroyed before the explosive with detonator is initiated. This has been proved by field measurements which indicate that the misfire is up to 24% when delay time is equal to or over 500 ms between holes, but the misfire in similar blasting conditions is less than 18% when delay time is 100 ms. In brief, a too-long delay time in production blasting should be avoided. 5. Low quality of explosives and incorrect charge operation. Sometimes it happens that the explosive in a borehole does not detonate well because either the quality of the explosive is low or the charging operation of the explosive is incorrect. For example, some large air gaps are left in the blastholes due to too-fast feeding. 6. Too-large initiation errors for the neighboring holes. For example, the hole planned to initiate earlier is actually initiated after the hole planned to initiate later. This may occur when pyrotechnic detonators are used. 7. Connection between detonator wires and detonating cord. For example, when a double knot is not used to connect a detonator wire and a detonating cord, the detonator may not be initiated, as described in the chapter: Explosives and Detonators. 8. Bad detonators. This seldom happens but it does sometimes occur. 9. Incorrect detonator position. As analyzed in the chapter: Primer Placement, in normal blast condition, it should be avoided that a detonator is placed at or close to the collar of the borehole. 10. Simultaneous initiation of two or more neighboring blastholes. In the Ridgeway mine, two middle holes would be first initiated at the same time in a ring, and then their neighboring holes should follow. But the second delay often had a malfunction [11]. This initiation often results in the malfunction to the neighboring holes since such a simultaneous initiation produces a much larger fragmentation radius Rf than a single hole initiation. 11. Borehole damage due to the variation of in situ stress field. According to borehole filming investigation, the borehole damage includes various forms such as deformed borehole, sheared borehole, broken-out borehole, and jammed borehole [12].
19.10 FRAGMENTATION, ORE RECOVERY, AND MINING PROFIT As mentioned previously, the main disadvantages of sublevel caving are high ore loss and high dilution, and both can be up to 20% or even more in large-scale sublevel caving mines. If an iron ore mining company has an annual ore production of 30 million tons with a grade of 64%, the total ore loss per year will reach 6 million tons of iron ore. What a figure for ore loss! We will see that rock fragmentation can be largely improved by blasting, and a better fragmentation can increase both productivity and ore recovery and finally raise the profit of a mining enterprise. These subjects will be described in the chapter: Effect of Blasting on Engineering Economy.
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19.11 SAFETY, ENVIRONMENT, AND VIBRATION CONTROL 19.11.1 Safety Working safety in underground mining is related to various factors such as blast operation, rock conditions, and mining methods. Here we do not discuss the accidents such as fires, traffic casualties, ventilation, water, electricity problems, and so on. Only blast-related safety problems are accounted for here, and some of them are roof problems, rock spalling, rock bursts, rock falls, and brow damage. Excluding brow damage, which has been discussed in this chapter, the others will be described in the chapter: Safety in Rock Engineering.
19.11.2 Environment It is a common problem that rock blasting in underground mines makes local underground water contaminated more or less. There are mainly two reasons for the problem. (1) In multihole blasting, the explosive in one or more holes is not detonated. The undetonated explosive will partly go into the local underground water system, and partly be transported to mineral processing plants together with ores. (2) When an upward hole is charged, for instance in a sublevel ring, the explosive sinks in boreholes and then falls to the floor during and after charging, as shown in Figs. 19.15 and 19.16, which frequently happens in production charging, especially when emulsion is pumped into a wet hole. The main reasons for explosive sinking are (1) great borehole deformation due to the variation of the local stress field, (2) water in boreholes, and (3) low stickiness of the explosive. Therefore, in order to avoid explosives falling out of b oreholes,
FIGURE 19.15 Precharged explosive sank to collars. (a) Upper parts (over 15 m from collars) of boreholes in sublevel caving ring were precharged; after a few weeks the explosive sank to the collars; (b) one borehole from the ring. The pictures show the roof of the production drift.
FIGURE 19.16 Explosive fell to the drift floor after it was loaded into the blastholes.
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precharge should not be used in the cases where boreholes are either wet or in a large deformed rock mass. In a ddition, to reduce incidents of falling explosives from precharged boreholes, stickiness of an explosive should be increased.
19.11.3 Ground Vibrations The blast-caused ground vibrations will be described in the chapter: Reduction of Ground Vibrations.
19.12 SUBLEVEL-CAVING BLASTING IN THE FUTURE The mining industry has the following missions: l
To increase ore recovery or reduce resources loss in mining operation; To increase profit in the mining process; l To secure safety in underground mining. l
In order to complete these missions, the following are important measures: l
l l l l l l l l l l l
Correct primer placement; Correct stemming; Proper charge diameter to meet the requirement for sufficient VOD; Use of dividing single blasting (DSB) method if necessary; Appropriate control of ore flow such as by leaving an ore layer as a buffer; High specific charge or smaller burden; Double-ring blasting; Efficient open cut and drifting; Correct mining layout and mining sequence; Increase in ore recovery from previously finished sublevels and some single rings where ore extraction is going on; Interfering high-stress state or releasing high stresses in the high-stress regions; Reducing borehole damage.
Some of these measures were described in the previous chapters such as primer placement, but others are not well developed. For example, the mature techniques on interfering high-stress state and reducing borehole damage are few. This is because sublevel caving mining meets with more and more challenges as mining depth increases. Some of the challenges are: l
High-stress state in deep mining. With increasing mining depth, the stress state becomes higher and more severe. The high-stress state can result in different problems, some of which are: (a) large borehole deformation which causes the precharged explosive to be pushed out of the borehole; (b) rock bursts or seismic events, especially when a fault is involved; the rock bursts or seismic events can seriously damage production drifts, including boreholes; (c) various problems of borehole damage [12] (see Fig. 19.17). These problems largely undermine rock blasting and adversely affect fragmentation and ore recovery. l Effect of hanging wall collapse on borehole stability. Each time when a hanging wall suddenly collapses or caves in, a great amount of energy in the form of stress waves is released, partly into the solid rock mass and partly into the previously caved waste rocks. In consequence, the stress field and stress state in the mine will be affected. The rapid change in the stress field and stress state can result in borehole deformation and damage. l Effect of production drilling on borehole stability. Production drilling is often done a long time, for example, one year, before rock blasting. This means that boreholes are loaded a long time before they are charged. Therefore, when this time is shortened, the borehole deformation and damage may decrease. This is to be confirmed by further study. l Effect of blasting on muckpile’s compaction and ore recovery. Fig. 19.18 shows a hanging-up and rock–ore sandwiches in sublevel caving mining. Probably, a simultaneous blasting of two or three middle holes is a major reason for the hanging-up and the compaction. This compaction adversely affects ore recovery, and the hanging-up has a negative effect on production and safety. l Optimum fragmentation from mine to mill. Optimum fragmentation has a great economic potential in underground mining. It may be realized by two ways. One is to increase energy efficiency in each operation such as drilling, blasting, crushing, and grinding. For example, to increase energy efficiency in blasting by means of efficient stress superposition. The other way is to increase the energy input to the blasting operation, so that the energy consumption in downstream operations such as crushing and grinding is dramatically reduced. A detailed description of optimum fragmentation will be presented in the chapter: Optimum Fragmentation.
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FIGURE 19.17 Borehole instability problems in sublevel caving. (a) Production blasthole is sheared, broken, and jammed; (b) production blasthole is fractured.
FIGURE 19.18 Hanging-up and compact ore–rock sandwiches in sublevel caving. The middle part of the picture is the drift. The place at which the lamplight points is a layer of waste rock. Over and below this layer is a layer of ore left. Another layer of waste rock can be seen close to the upper part of the drift.
19.13 CONCLUDING REMARKS 19.13.1 Characteristics of Sublevel Caving Blasting The main disadvantages of sublevel caving are high ore loss and high dilution, and both ore loss and dilution can be up to 20% or even more. However, ore loss can be reduced by improving rock blasting. The spacing varies to a great extent, which results in an uneven distribution of explosive within the ore mass to be blasted. Accordingly, spacing, charge length, and initiation plan are to be determined carefully.
19.13.2 Subdrilling, Deviation, and Inclination As in open pit blasting, subdrilling in sublevel caving is also necessary for both a narrow ore body and the rings directly under hanging walls. This is often overlooked. In the case of a large deviation in rock drilling, delay time and initiation sequence have to be adjusted so as to consider the actual borehole position in the final blast operation. An inclined ring is better for brow protection than a vertical ring.
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19.13.3 Single-Ring Blasting and Double-Ring Blasting For a narrow ore body, it is better to choose single-ring blasting instead of double-ring blasting so as to reduce permanent ore loss. For a large ore body, double-ring blasting with an effective extraction control should be favorable for fragmentation and ore recovery.
19.13.4 Delay Time The principles for determining delay time described in the chapter: Delay Times are valid for sublevel caving blasting. In order to obtain optimum delay time, either detonation wave (or borehole pressure) or stress wave close to the blasthole is necessary. At the same time, electronic detonators are recommended.
19.13.5 Brow Damage Serious brow damage not only gives rise to lower ore recovery and higher dilution, but also undermines safety for the people working in the field. Brow damage can be reduced by taking various measures such as multiprimer placement in a blasthole, correct detonator position, reduction of specific charge, high-VOD explosive, proper initiation plan, and so on.
19.13.6 Misfire Misfire is still a serious problem in sublevel caving and other types of underground mining. The main causes of misfire are pressure transients, too-small spacing, sympathetic detonation due to fractures or empty rooms within the rock mass between two holes, too-long delay time between neighboring holes, low quality of explosives, incorrect charge operation, wrong detonator position, and so on.
19.13.7 Other Issues Related to Sublevel Caving Blasting There are other important issues related to sublevel caving blasting such as open cut, drifting, stemming, air deck, detonator placement, fragmentation, mining cost, mining safety, and environmental influence. Some of these have been described previously, and the others will be discussed later in the book.
19.13.8 Underground Blasting in the Future Underground blasting in the future will have to pay attention to problems caused by a high-stress state, such as borehole instability, high confining pressure in blasting, effect of blasting on muckpile and ore flow, and optimum fragmentation.
19.14 EXERCISES 1. List the advantages and disadvantages of the sublevel caving method. If you are assigned to a project aiming to reduce the ore loss in an underground mine where sublevel caving is used, do you know which methods can be used to reduce the ore loss in sublevel caving? 2. In Fig. 19.1, if blasthole 5 is not detonated, which consequences can be caused for mining production and environmental protection? 3. In Fig. 19.2 the subdrilling is necessary. If there is no subdrilling in the rings under the hanging wall, what consequence may be caused for the mining production? 4. In the cases of deviation shown in Fig. 19.4c and d, how do you make your blast plan in order to ensure that all of the holes are detonated? 5. In which case may a double-ring blast be better for fragmentation and ore recovery? To ensure that the recovery is high in a double-ring blast, what should you do? 6. List the measures for reducing brow damage in sublevel caving. Explain each measure by using your own words. 7. Use the stress wave theory in the chapter: Stress Waves to explain why there is much spalling in the roofs shown in Fig. 19.14a. 8. In normal conditions, the emulsion explosive pumped into a blasthole can well stay in the hole for a long time (up to two months) without sinking. But sometimes or in some mining areas, the explosive charge markedly sinks, as shown in Fig. 19.15. Give the possible reasons for the sinking. 9. Use Eqs. 19.1 and 19.2 to prove this relation: Bebv > Bebi.
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REFERENCES [1] Hustrulid W, Kvapil R. Sublevel caving—past and future. Newsletter—Australian Centre for Geomechanics 2009;33:1–6. [2] Janelid I, Kvapil R. Sublevel caving. Int J Rock Mech Min Sci 1966;3:129–53. [3] Zhang ZX. Effect of double-primer placement on rock fracture and ore recovery. Int J Rock Mech Min Sci 2014;71:208–16. [4] Ren FY. Ore flow theory. Beijing: Press of Metallurgical Industry of China; 1994. [5] Zhang ZX, Naarttijärvi T. Reducing ground vibrations caused by underground blasts in LKAB Malmberget mine. Int J Blast Frag 2005;9(2):61–78. [6] Olsson M, Nyberg U, Fjellborg S. Controlled fragmentation in sublevel caving-first tests. Sweberec Repport 2009:2.(in Swedish). [7] Zhang ZX. Reducing eyebrow break caused by rock blasting in Malmberget mine. In: Proceedings of thirty seventh ISEE annual conference, San Diego, USA, 6–9 Feb., 2011, [8] Dawes JJ, McKenzie CK, Liddy TJ. Interaction between blast design variables: experimental and modelling studies. In: Proceedings of first international symposium on rock fragmentation by blasting, Luleå, Sweden, August, 1983, Vol. 1, p. 265–287. [9] Nyberg U, Fjellborg S. Controlled drifting and estimation of blast damage. In: Proceedings of first world conference on explosives and blasting technique, Munich, Germany, 6–8 September, 2000. Rotterdam: Balkema, p. 207–216. [10] Mencacci S, Farnfield R. The measurement and analysis of near-field pressure transients in production blasting. In: Proceedings of second world conference on explosives and blasting technique. Lisse: Swets & Zeitlinger; 2003. p. 467–473. [11] Brunton ID, Chitombo GP. Modelling the impact of sublevel caving blast design and performance on material recovery. In: Proceedings of ninth international symposium on rock fragmentation by blasting. London: Taylor & Francis Group; 2009. p. 353–362. [12] Ghosh R, Zhang ZX, Nyberg U. Borehole instability in Malmberget underground mine. Rock Mech Rock Eng 2015;48:1731–6.
Rock Blasting in Underground Mining Chapter Outline 19.1 19.2
19.3
19.4
19.5
19.6 19.7 19.8
Advantages and Disadvantages of Sublevel Caving Drilling and Charging Plan 19.2.1 Burden and Spacing 19.2.2 Length of Blastholes 19.2.3 Uncharged Length 19.2.4 Precision of Rock Drilling Inclination of Rings 19.3.1 Loading and Ore Flow 19.3.2 Brow Protection Single- or Multiple-Ring Blasting 19.4.1 Single-Ring Blasting 19.4.2 Double-Ring Blasting Delay Time and Initiation Sequence 19.5.1 Delay Time 19.5.2 Initiation Sequence 19.5.3 Initiation and Delay Time in Special Cases Open Cut and Drifting Stemming, Air Deck, and Detonator Placement Back Break and Brow Damage 19.8.1 Back Break 19.8.2 Brow Damage
373 374 374 375 377 377 377 377 379 380 380 381 382 382 382 382 383 383 384 384 385
19.9 Misfires 19.10 Fragmentation, Ore Recovery, and Mining Profit 19.11 Safety, Environment, and Vibration Control 19.11.1 Safety 19.11.2 Environment 19.11.3 Ground Vibrations 19.12 Sublevel-Caving Blasting in the Future 19.13 Concluding Remarks 19.13.1 Characteristics of Sublevel Caving Blasting 19.13.2 Subdrilling, Deviation, and Inclination 19.13.3 Single-Ring Blasting and Double-Ring Blasting 19.13.4 Delay Time 19.13.5 Brow Damage 19.13.6 Misfire 19.13.7 Other Issues Related to Sublevel Caving Blasting 19.13.8 Underground Blasting in the Future 19.14 Exercises References
387 387 388 388 388 389 389 390 390 390 391 391 391 391 391 391 391 392
There are clear differences between production blasting in underground mines and that in open pit mines. First, in many underground mines a production blast consists of a fan-shaped blastholes in each row. Such a row is called a fan or a ring, which appears in different mining methods such as sublevel caving, block caving, and open stopping. The blastholes in a ring are not parallel to one another. Second, in most underground mines production blastholes are often drilled upward instead of downward as in open pit mines. The upward blastholes make it difficult to apply stemming. Because the sublevel caving method is widely used in various types of underground mines, we will focus on the blasting in sublevel caving in this chapter. However, the principles in sublevel caving blasting are applicable to other underground mining methods. Rock blasting in sublevel caving deals with drilling plan, charging plan, quantity of rings in each blast, back break, brow damage, and misfires. These will be described in detail in this chapter. In addition, we will outline delay time, explosive initiation, open cut, drifting, stemming, detonator placement, relation between fragmentation and ore recovery, and effects of blasting on safety and the environment since they are described in detail in other chapters. Last, blasting in sublevel caving in the future will be discussed.
19.1 ADVANTAGES AND DISADVANTAGES OF SUBLEVEL CAVING Sublevel caving is a widely used mining method in metallic mines all over the world. The modern sublevel caving was possibly developed in the iron mines of Sweden [1]. This method has many advantages with regard to safety and mechanization [2]. Because all of mining operations are carried out only in drifts, the safety in sublevel caving is relatively good, compared with other mining methods such as cut and fill, room and pillar, and so on. In addition, since big machines can be used in sublevel caving and mechanization and automation are available, mining production can be performed on a very Rock Fracture and Blasting: Theory and Applications Copyright © 2016 Elsevier Inc. All rights reserved.
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large scale. For example, automation or remote control in production drilling, underground transportation, extraction, and the like has been realized for decades in a number of large mines such as Kiruna and Malmberget mines. However, sublevel caving has some disadvantages. High ore loss and high dilution are two of them. It is not unusual for a sublevel caving mine to have ore loss up to 20% in the mining process. In brief, the major advantages and disadvantages of sublevel caving can be concluded as follows. The major advantages are: l
Good safety; High mechanization and automation; l Large-scale production and high productivity. l
The major disadvantages are: l
High ore loss; High dilution.
l
Therefore, how to eliminate these disadvantages is a challenging task for the sublevel caving method. Fortunately, production tests have shown that ore loss can be markedly reduced by improving blasting. This will be presented in the chapter: Effect of Blasting on Engineering Economy.
19.2 DRILLING AND CHARGING PLAN 19.2.1 Burden and Spacing 19.2.1.1 Burden How to determine burden in rock blasting has been described in the chapter: Burden and Spacing. Burden is dependent on more factors in underground mining than in open pit mining. For example, burden is related to mining scale, which depends on the sizes of equipment such as loading and drilling machines; burden influences ore recovery in sublevel caving, too. For instance, if burden is too large, it is difficult for a loading machine to reach the ore in the farthest muckpile. This makes part of the ore in the muckpile remained. In this case, if the ore body is narrow or only one drift is on each level, the remained ore (or at least part of it) will become a permanent loss. In addition, burden affects vibrations. In brief, the following factors are related to burden: l
l l l l
Production scale; Sizes of drilling, charging, and loading machines; Fragmentation; Ore recovery; Ground vibrations.
In blast design of sublevel caving, it is suggested that the burden should be determined on the basis of the analysis in the chapter: Burden and Spacing combined with empirical parameters. In present sublevel caving, the largest borehole is about 115 mm in diameter, and corresponding burden is often 3 m or 3.5 m in some mines. These parameters may be taken as a reference but not necessarily as the best choice since no study indicates they are optimum.
19.2.1.2 Spacing Drilling and blast plans in sublevel caving are different from those in open pit blasts. As shown in Fig. 19.1, a ring in sublevel caving looks like a fan. Note that the spacing in the sublevel ring is not a constant but the spacing in open pit blasting always is. From Fig. 19.1 we see that in the upper part of the ring the spacing is much larger than it is in the lower part; that is, the spacing for two neighboring holes varies from their collars to bottoms (eg, the collar of borehole 4 is at C). In addition, the boreholes all have different lengths and different inclinations. If only one kind of explosive is well distributed in a borehole, that is, the explosive and its density do not vary along the borehole, the varying spacing has a negative impact on fragmentation since the upper part of the ring always has a lower specific charge. This makes the ore in the upper part poorly fragmented, compared with the ore in the lower part. Consequently, in sublevel caving it is often the case that as soon as extraction starts after a ring is blasted, the sizes of fragments increase with increasing ore extraction, meaning that boulders usually come from the upper part of the ring. To reduce boulders or improve fragmentation, the spacing in the upper part must be reduced. In this way, the uncharged length in the lower part of every next hole must be increased in order to avoid misfires. Two other alternatives are that (1) in the upper part of ring a stronger explosive with higher velocity of
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FIGURE 19.1 Production ring in sublevel caving. Rf is the radius of the fragmentation region of each blasthole; Sch is the distance from the end of the explosive charge in one hole to another that is nearest. There should be Sch ≥ Rf if delay time is not very short. The uncharged part of hole 6 is too short. The correct position should be at the other end of the short line at I.
detonation (VOD) is charged, and (2) one ring is divided into two parts—upper part and lower part—in blasting and a very short delay time is applied in the upper part to make full use of stress wave superposition [3]. Although the first alternative is difficult to implement in production today, it will probably be more possible in the future. An introduction to the second alternative will be made in the chapter: Reduction of Ground Vibrations. Up to now boulders have been a big problem in underground mining. On one hand, the current level of specific charge, which is often 1 kg/m3 more or less, is too small in underground mining, indicating that the specific charge may have a great space to increase from the present level. On the other hand, the blasting methods in current production blasting have much to improve. For example, detonator placement, stemming, delay time, and the like all need a scientific design.
19.2.2 Length of Blastholes As soon as the sublevel height is determined, the size of a ring and the length of blastholes in a ring can be decided. We discuss two cases: one for a large ore body and the other for a narrow ore body.
19.2.2.1 Large Ore Body In a large ore body, there are two types of rings: (1) the rings under the hanging wall that is composed of waste rock mass, as shown in Fig. 19.2a, and (2) ordinary rings as shown in Fig. 19.2c. In ordinary rings, there must be a zone called undrilling or buffer left in the top of the ring, as shown in Fig. 19.2c. In other words, a length indicated by “Undrilling” is left without drilling. If the boreholes in a ring reach its upper boundary, after production in the upper level is finished, some fragments
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FIGURE 19.2 Subdrilling and undrilling in wide ore body. (a) Rings under hanging wall; (b) large ore body with hanging wall; (c) ordinary rings.
may fall into and block the boreholes. This will make it impossible to charge explosive in the upper part of the boreholes. As a result, fragmentation will be worse. Therefore, we need the undrilled part. In sublevel caving 1 m or so can be tried for the length of undrilling. If it is too long, the fragmentation will be worse. If the ring is under a solid hanging wall, as shown in Fig. 19.2a and b, the subdrilling is necessary, like in open pit drilling. Otherwise, without subdrilling, that is, the bottoms of the boreholes are at the upper boundary of the ring after blasting, the ore mass surrounding the bottoms will not be completely broken since confinement is strong there. This is similar to blasting in open pit mines. As a consequence, that part of ore mass will be left as a permanent loss. In order to avoid such an ore loss, a subdrilling is necessary. It is suggested that 0.5–1.0 m can be taken as a reference for the size of subdrilling in sublevel caving.
19.2.2.2 Narrow Ore Body In a narrow ore body, there is often one drift on each production level, as shown in Fig. 19.3. Several boreholes close to a hanging wall must have subdrilling so as to avoid ore loss near the boundary between ore body and hanging wall. In contrast, without subdrilling, the ore mass close to the boundary will not be well broken due to high confinement surrounding
FIGURE 19.3 Sublevel ring in a narrow ore body.
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the bottoms of boreholes. Finally, this ore mass will be left as a permanent loss. Such a permanent ore loss may be considerable if a mine has many narrow ore bodies and its production is large.
19.2.3 Uncharged Length In a ring blast, the length of the uncharged part of a borehole must be carefully determined due to the varying spacing. Such a length is mainly dependent on the fragmentation radius Rf of each hole; see Fig. 19.1. There are two different situations: (1) as delay time is longer than the duration of fragmentation propagation between the two holes, and (2) as delay time is shorter than the duration. These are to be discussed in the following.
19.2.3.1 As Delay Time Is Longer Than Fragmentation Time Under this condition, as shown in Fig. 19.1, if the distance Sch is larger than the fragmentation radius Rf of the first borehole, detonation of the first-initiated hole will not destroy the second-initiated one. In contrast, if the distance Sch is smaller than Rf, the detonation of the first-initiated hole will destroy the blasthole to be initiated immediately following the first, and consequently the explosive in the latter (second) will not detonate. As a conclusion, when delay time is longer than the fragmentation time between two neighboring holes, the distance Sch must be equal to or larger than Rf. If so, the detonation from one borehole will not influence or destroy its neighboring holes. According to the experience of LKAB, the fragmentation radius Rf is approximately 1.5 m when the borehole diameter is 115 mm and emulsion explosive is used. For other mines, it is better to determine Rf by carrying out their own tests.
19.2.3.2 As Delay Time Is Shorter Than Fragmentation Time In this case, the distance Sch can be shorter than the fragmentation radius Rf since the second-initiated hole will start to detonate before fragmentation from the first-initiated hole reaches the second hole. This will happen when electronic detonators with a short delay time are used. In this case, efficient stress wave superposition from neighboring holes may be realized. This case is an attractive area for production blasting in both underground and surface blasting in the future. In order to achieve a successful blast result, it is necessary to measure detonation wave or stress waves in the rock. Based on this, a correct delay time can be determined. Otherwise, if the determination of delay time is not based on detonation wave measurement and stress wave analysis, the delay time that is chosen may not be correct. As a consequence, the blast result will not be very good.
19.2.4 Precision of Rock Drilling The deviation of rock drilling in sublevel caving can be divided into three types, as shown in Fig. 19.4. (1) The borehole curves in two directions from its original central line, as shown in Fig. 19.4a and b. (2) The angle of the borehole changes but the collar position does not vary and the borehole does not curve, as shown in Fig. 19.4c. (3) The collar position is changed and the borehole is curved in only one direction, as shown in Fig. 19.4d. Any type of deviation shown in Fig.19.4 makes rock fragmentation worse. For example, two extremes of deviation in Fig. 19.4a are (1) the bottom distance between two holes 4 and 5 is much larger than that planned, which may produce boulders there; (2) two holes 5 and 6 are closer to each other, meaning that either sympathetic detonation occurs if they are too close, or one is destroyed when the other is detonated. The latter will also happen if the deviation is like in Fig. 19.4c. This will occur when pyrotechnic detonators are used since their delay time between two adjacent holes is long. The similar misfire will happen if the deviation in Fig. 19.4d appears. Therefore, precise drilling is important in sublevel caving as well as in other similar blasting.
19.3 INCLINATION OF RINGS There are usually two types of ring inclinations: (1) inclined rings as shown in Fig. 19.5 and (2) vertical rings as shown in Fig. 19.6. We will see the differences between the two cases. Assuming the sublevel height, drift sizes, burden, borehole diameters, length of stemming (or uncharged length), explosives, and other parameters are all identical for the two cases, we will compare the two different inclinations in the following.
19.3.1 Loading and Ore Flow If the widths of the muckpiles in both cases are equal, as represented by Wf in Figs. 19.5 and 19.6, the maximum distance available for loading along the drift axis will be the same for the two cases. In addition, the contact areas between waste
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FIGURE 19.4 Different types of deviation. (a) Deviation of borehole 5 occurs in the plane of the paper, while the deviation of other holes happen in the plane perpendicular to the paper plane. (b) Cross section of (a). (c) The collar position of hole 3 is not changed but its angle is altered. (d) The collar position of hole 5 is changed and the hole is deviated.
FIGURE 19.5 Inclined ring.
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FIGURE 19.6 Vertical ring.
rock and fragmented ore on the top of the rings are equal in the two cases, meaning that ore dilution caused by the mixture of waste rock and ore from the tops of the rings will be the same. If the burden in both cases is constant in the horizontal direction, the inclined ring and the vertical one will produce the same amount of ore, as shown in Figs. 19.5 and 19.6. For ore flow, it is difficult to find a marked difference between the two cases.
19.3.2 Brow Protection For brow protection, the inclined ring is better than the vertical ring, according to the following analysis. Under otherwise identical conditions, we assume that detonation direction is up–down and the velocity of detonation is greater than the velocity of the P-wave in the ore or rock mass. In this case, when the last part of the explosive charge above stemming is exploded, the front of the P-wave from blasting will first reach both the front surface and the roof of the drift, as shown in Fig. 19.5. After the P-wave arrives at the surfaces, two tensile waves will be caused by the reflection of the P-wave. Tensile rock fracture will start within the region surrounded by the front of the tensile wave (notice that there is also a compressive P-wave there, but the resultant stresses will mostly be tensile). Therefore, we can further assume that the size of brow damage is proportional to the size of the tensile wave region. In other words, the larger the tensile wave region is, the larger is the region of brow damage. As shown in Figs. 19.5 and 19.6, after the same time when detonation at the point O occurs, the front of the P-wave arrives at the circle indicated by the front of the compressive P-wave. Let the radius of this circle be R. In the inclined ring we assume that the brow damage Bebi is equal to the distance from G to H (in reality Bebi must be smaller than GH), and the length Lst of stemming is the distance between O and G. Since the angle of the inclined ring is bb, we can find: LFG = Lst cos β b LOF = Lst sin β b where LFG is the length of FG and LOF is the length of OF. Then we may get that the brow damage Bebi in the inclined ring is equal to
Bebi = R 2 − ( Lst sin β b )2 − Lst cos β b
(19.1)
In the vertical ring we can easily see that the corresponding brow damage Bebv is equal to
Bebv = R 2 − L2st
(19.2)
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By comparing Eqs. 19.1 and 19.2, we can find that as long as bb < 90 degrees, we have Bebv > Bebi
(19.3)
This formula indicates that the vertical ring produces larger brow damage than the inclined ring.
19.4 SINGLE- OR MULTIPLE-RING BLASTING In sublevel caving, one blast often contains one ring, but in some mines one blast consists of two rings. What are the differences between the two methods, and which one is better? This will be discussed in the following.
19.4.1 Single-Ring Blasting Before a production ring is blasted, the front face of the ring is always against waste rock (often boulders) with a mixture of the ore fragments remained from previous blasts. In a normal condition, the waste rock as well as the ore fragments remained can freely flow, meaning that the front face is a partial free surface. Therefore, a compressive stress wave can partly go through the face into the caved waste rock. The major advantage of the single-ring blast method is that the blasted ore can be extracted as much as possible. To make the analysis simple, assume that the farthest position which the bucket of a loading machine can reach is F in Fig. 19.7. Notice that OF is a vertical line. At the same time, assume that the rest angle of flowing ore is brest. Thus, the ore within the area E–G–F–H–E cannot be loaded out, and it will be remained. In a narrow ore body, that ore may partly become a permanent ore loss. As shown in Fig. 19.7, when drift height is Hdrift(= OF) and the width of ore flow is Wflow(= HO), we have H drift = tan β rest Wflow This equation can be used to estimate the amount of the ore remained. When Hdrift = 5.5 m and brest = 60 degrees, we get Wflow = 3.17 m. As shown in Fig. 19.8b, some ore close to the foot wall may be left if the angle bfoot of the foot wall is smaller than the angle brest of the flowing ore. If both bfoot and brest are known, the amount of ore left can be calculated. This part of the ore is almost impossible to recover from a lower production level (level 2 in Fig. 19.8b). In order to reduce such ore loss, the drilling plan should be made carefully, for example, to let the distance between the last hole (ie, first hole from the right side of
FIGURE 19.7 Ore extraction in single-ring blasting. Ring Ri has been blasted but ore extraction has not started yet. Ring Rj is to be blasted next time, and its front face is at AO, which is a partial free surface either during or after extraction.
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FIGURE 19.8 Ore loss. (a) Loss in large and wide ore body; (b) loss in small and narrow ore body.
Fig. 19.8b) and the foot wall be as small as possible. In addition, in the loading operation, the loading machine may load only the ore near the foot wall at the drift in the beginning of loading. This can make the ore close to the foot wall flow down easily. All in all, for a narrow ore body as shown in Fig. 19.8b, it is better to use the single-ring blasting method in order to reduce permanent ore loss.
19.4.2 Double-Ring Blasting As for two rings that are blasted together in one blast, we discuss two cases: one for a narrow ore body and the other for a large ore body. In a narrow ore body as shown in Fig. 19.8b where there is only one drift on each production level, the remained ore shown by the area E–G–F–H–E in Fig. 19.9 is much more than that in Fig. 19.7 where one ring is blasted. As
FIGURE 19.9 Ore extraction in double-ring blasting. Two rings R1 and R2 are blasted together in one blast.
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discussed previously, if a narrow ore body is not inclined very steeply, the ore remained in the narrow ore body cannot be wholly recovered in the lower level. Therefore, the double-ring blasting method is not a good choice for narrow ore bodies. In a large ore body as indicated in Fig. 19.8a, the situation is different from that in a narrow ore body. The ore left in most rings, excluding the rings near the foot wall, on a production level (eg, Level 1 in Fig. 19.8a) can be wholly extracted in the lower levels, theoretically. Thus, the ore loss from double-ring blasting will not be larger than that from single-ring blasting. In addition to ore loss, we will talk about fragmentation and dilution in double-ring blasting. On fragmentation, for double-ring blasting, the throw of fragments from the second ring is largely hindered by the first ring, making a collision happen between the fragments from the first and the second rings and resulting in secondary fragmentation of these fragments. As a result, the fragmentation in the second ring is often better than that in the first ring or the single ring. This is consistent with the field observations and production tests in open pit blasting introduced in the chapter: Rock Blasting in Open Pit Mining. Since fragmentation in a double-ring blast is generally better than that in a single-ring blast, the dilution in the doublering blast may be reduced through controlling ore extraction. For example, the low ore loss and low dilution method tested in China [4] is to leave a certain amount of ore in each production level so as to always form a layer of ore over the ongoing production level. By doing this, the interface area between the ore and the waste rock can be greatly reduced, and finally the dilution will decrease. This method is considered to be efficient in large ore bodies. Therefore, the double-ring blasting method together with an effective control of ore extraction is worthy of testing in sublevel caving with large ore bodies.
19.5 DELAY TIME AND INITIATION SEQUENCE 19.5.1 Delay Time The principles for determining delay time have been described in the chapter: Delay Times, and they are valid for sublevel caving blasting. As described in that chapter, in order to obtain optimum delay time we have to know the stress waves in the rock. Up to now, the applications of electronic detonators have not shown a much better result such as improved fragmentation in some mines. One of the major reasons is that the delay time used is out of efficient wave superposition. As soon as the stress wave (or detonation wave, which can be taken as a reference) is known, the principles in the chapter: Delay Times can be used to determine a correct delay time.
19.5.2 Initiation Sequence In normal conditions of sublevel caving, the initiation in a ring blast starts at the middle blasthole such as No. 4 in Fig. 19.10 since it is thought that the confining pressure in front of the middle hole is smaller than that of others. However, in some cases, for example, as ground vibrations caused by blasting must be controlled, the initiation sequence needs to be changed, rather than just starting at the middle hole. That is to say, if needed, the initiation may start from any hole except for the first one (eg, No. 1) and the last one (eg, No. 7). This has been proved by the vibration control method used in the Malmberget mine, where the initiation starts from the blasthole nearest to either the first or the last hole [5]. Optimistically, the production blasts indicate that the new initiation sequence (starting from any holes except for the middle, first, and last ones) does not adversely affect rock fragmentation.
19.5.3 Initiation and Delay Time in Special Cases As mentioned previously, deviation is a common phenomenon in rock drilling. It is quite often that the deviation occurs in one or more boreholes in a ring, as shown in Fig. 19.10 where borehole 5 has a deviation. The deviation of borehole 5 increases the distance between 4 and 5 at their bottoms and decreases that between 5 and 6. In this case, if the initiation starts from borehole 4 and the delay time between two neighboring holes is normal (ie, not instantaneous or very short), the consequence will be (1) that boulders are produced in the upper ring between boreholes 4 and 5, and (2) that when borehole 5 is blasted, borehole 6 may be destroyed at the same time. This happens since the Sch between holes 5 and 6 is smaller than Rf. In order to avoid this, the initiation and delay time in the ring can be adjusted. For example, when the initiation sequence is not changed, the delay time between boreholes 5 and 6 should be very short or even simultaneous so that both holes are safely blasted.
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FIGURE 19.10 Initiation sequence in a sublevel ring. When hole 5 is fired prior to 6, the EF part of hole 6 will be destroyed since Sch < Rf.
19.6 OPEN CUT AND DRIFTING Open cut and drifting have been described in the chapter: Rock Blasting in Open Cut and Tunneling. How to make a plan for open cut and drifting in sublevel caving may follow the principles there.
19.7 STEMMING, AIR DECK, AND DETONATOR PLACEMENT Stemming, air deck, and detonator placement have been described in the chapters: Stemming and Change Length; Air Deck and Smooth Blasting; Primer Placement. Here we just want to address some points. First, stemming must be applied in underground mining, even though an ideal stemming is difficult to implement in underground mines. As described previously, both material and sizes of stemming affect blast results. Therefore, it is important in rock blasting to choose a proper material and correct sizes of stemming. Second, air deck and other charge techniques may be employed in underground mines if necessary. Third, if boreholes are short, one detonator or primer is enough to achieve a good result in blasting. In this case, the best detonator position is at the middle of the charged borehole (not the middle of the total length of the borehole) with a correct stemming. If boreholes are long and multiple-detonator positions are available, two detonators positioned in each hole may be well applied. If the two detonators at different places in a hole are initiated simultaneously, shock collision will happen and it will give rise to the final shock pressure being more than the sum of the two shocks. This will to a great extent strengthen rock fragmentation. This double-primer method has been tested in the Malmberget mine with a result of less
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brow damage and higher recovery [3]. Moreover, no damage to the structures nearby has been found due to this method. However, as the double-primer method is used in sublevel caving, ore flow may change markedly, depending on the delay time. Accordingly, a prestudy or a pretest is necessary so that a correct blast design is made finally.
19.8 BACK BREAK AND BROW DAMAGE 19.8.1 Back Break Back break and brow damage are common in sublevel caving, but they are sometimes the same thing. Therefore, a definition is needed for both. As shown in Fig. 19.11, part of the burden in ring R2 is broken off by the blast of R1. The average thickness of this part of the burden along the whole ring face is called back break, indicated by Bb. However, brow damage, indicated by Beb, means that part of the burden at the roof of drift is broken off. Brow damage is often larger than back break in sublevel caving. It is difficult to measure back break in the field. When a blasthole in a sublevel ring such as R1 in Fig. 19.11 is blasted, radial cracks will be produced surrounding the blasthole. Then when the compressive stress wave propagates to the front surface of the ring (ie, a partial free surface), it is reflected into a tensile wave. This tensile wave will travel toward the rings behind R1, such as R2 and R3, since the rock surrounding the blasthole needs enough time to be completely fractured. After the tensile wave passes R1, the rock behind it will start to fracture if the tensile stress is larger than the dynamic tensile strength of the rock. In this way, the rock behind R1 is broken; that is, a back break is formed. A back break is a typical spalling caused by a tensile stress wave as described in the chapter: Stress Waves. Remember that the compressive stress wave also contributes the back break, since it causes radial cracks surrounding the boreholes. To reduce the back break, the amplitude of the tensile stress waves must be reduced. This can be achieved by reducing the size of the blasthole. When doing this, one needs to ensure that fragmentation will not be worse. The factors influencing back break are sizes of blasthole, specific charge, VOD, primer placement, initiation plan, blast plan (single-ring or multiple-ring blasting), rock properties, and geological conditions. Other measures for reducing back break can be developed on the basis of the aforementioned factors. For example, by using a double-primer placement, brow damage has been reduced [3]. Probably, back break may be reduced by the same method. Back break in open pit blasting is quite similar to that in underground blasting, as measured by Olsson et al. [6].
FIGURE 19.11 Back break and brow damage.
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Since direct measurement of back break for whole blastholes in underground mining is difficult, we usually see the back break from the roof of drifts, as shown in the next section.
19.8.2 Brow Damage 19.8.2.1 Consequences Caused by Brow Damage Brow damage is also called brow break or eyebrow break in sublevel caving. Serious brow damage decreases ore recovery and increases dilution, and also brings about worse safety for the people working in the field. For example, miners must go to the caving area and stand under the broken brow. Brow damage is caused mainly by two parts of tensile stress waves: one from the front face of a blasting ring, and the other from the roof of the drift. We will make a detailed analysis in the following.
19.8.2.2 Factors Influencing Brow Damage Main factors affecting brow damage are detonator position, quantity of primers in a blasthole, specific charge, VOD, initiation plan, contact condition of the front face of the rings, and geological conditions of rock/ore.
19.8.2.3 Measures for Reducing Brow Damage The measures for reducing brow damage are correct detonator position, multiple-primer placement, decoupled charge, higher VOD, and stopping tensile stress waves such as by presplit blasting. In the following we will briefly introduce these measures. First, detonator position, as analyzed in the chapter: Primer Placement, largely influences brow damage in sublevel caving. Fig. 19.12 shows that when a detonator is placed at the middle of a charge, the brow damage is greatly reduced, compared with the damage when a detonator is placed near the collar. Second, on the double-primer method discussed in chapter: Primer Placement, as shown in Fig. 15.6 , when the explosive in the double-primer hole is completely detonated, the stress waves with all the energy are just distributed within the rock to be broken, while no wave comes to the brow area. From this point of view, the double-primer method is as good as placement of a single middle primer regarding brow protection. One more advantage of the double-primer method, compared with the single middle primer method, is that a shock collision happens, which results in fast rock fracture in the area between the two primers, as shown in Fig. 15.6. This fast fracture will mean there will be less of a tensile wave traveling to the brow area. In consequence, brow damage is reduced, as shown in Fig. 15.12, which indicates the result from electronic detonators [3]. As a matter of fact, brow damage can also be reduced by the double-primer method with Nonel detonators, as shown in Fig. 19.13.
FIGURE 19.12 Detonator position and brow damage. (a) When detonators were close to the collars, there was large brow damage and the burden of next ring (in charging process) was seriously broken off. The brow damage even extended to the ring behind the ring to be blasted next time; see the black area to the right of the dashed line. (b) When detonators were at the middle of charged holes, there was little brow damage and most of the burden of the next ring to be blasted was remained with no damage.
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FIGURE 19.13 Photographs of brow damage taken from two neighboring drifts in an ore body at Malmberget mine. An ordinary charge was in (a) with one primer at 15 m away from the collar, and a double-primer charge was in (b). The detonators were Nonel.
Third, we will show how to reduce brow damage by a decoupled charge (ie, using a cartridge charge). This method has been successfully tested in the Malmberget mine [7]. The principle of the method is simple: to reduce the tensile stress at and near the brow. In order to implement this, some holes, like No. 3 and No. 5, are charged with cartridge explosive (ie, a decoupled charge), as shown in Fig. 19.14. In this way, the amplitude of the tensile stress waves in the brow area decreases, and finally the brow damage can be reduced. The tests in the Malmberget mine indicated that the brow damage had been reduced from 3.0 to 2.3 m by using a cartridge charge [7]. Fourth, if the VOD of the explosive is very high, for example, it is much higher than the sonic velocity of the rock, the stress distribution surrounding the borehole will soon become more even. This may partly reduce the tensile stress wave coming to the brow area. Fifth, presplit blasting can be made between two rings in order to reduce brow damage. However, notice that this may markedly increase the cost and working hours for blast operation. In addition to these five measures, there are possibly other methods for reducing brow damage. This needs further study. An interesting result is that the ore recovery from both the single middle primer rings and the double-primer rings is higher than that from the ordinary rings according to the tests in the Malmberget mine [3,7].
FIGURE 19.14 Cartridge charge in the lower parts of boreholes.
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19.9 MISFIRES Malfunction or misfire is quite common in underground blasting. According to field measurements in open stopping operations, the misfires were up to 30% [8]; in drifting, the misfires were about 23% [9]; in sublevel caving production blasting, the misfires were from 10 to 18% in the Malmberget mine. The lower percentage was from the middle primer rings and the higher one from the lowest primer rings. Anyway, misfire is still a serious problem in underground mining. There are a number of factors causing malfunction or misfire: 1. Pressure transients. It is widely suspected that pressure transients, generated by detonating explosives and transmitted through the surrounding rock or stemming material, can cause significant malfunctions in both initiation systems and explosives [10]. Mencacci and Farnfield showed a picture of two detonators that were severely damaged by pressure transient in a multidecked blasthole [10]. This may happen in the production blastholes that are precharged with explosives and detonators. In other cases where the P-wave speed of rock is greater than the VOD of the explosive, malfunction of the explosive may occur since the explosive and detonator in front of the shock waves may be highly compressed by the stress waves from the rock before the shock waves from detonation arrive. 2. Too-short distance between two neighboring blastholes; that is, the distance between the two neighboring holes is smaller than the maximum fragmentation radius of the earlier detonated hole, and the delay time between the two holes is longer than the critical time during which rock fragmentation from the first-initiated hole can reach the lately initiated hole. This case often occurs when detonators are placed near the collars of blastholes according to the field measurements in sublevel caving [5]. 3. Sympathetic initiation of neighboring holes due to fractures or cavities within rock mass between the holes. In this case the detonation waves can directly travel from the detonating hole to its neighbor and initiate the explosive there. 4. Too-long delay time between two neighboring holes. A very long delay time makes cracks from a detonating hole propagate longer. If a number of such cracks reach the neighboring hole, then the latter will be destroyed before the explosive with detonator is initiated. This has been proved by field measurements which indicate that the misfire is up to 24% when delay time is equal to or over 500 ms between holes, but the misfire in similar blasting conditions is less than 18% when delay time is 100 ms. In brief, a too-long delay time in production blasting should be avoided. 5. Low quality of explosives and incorrect charge operation. Sometimes it happens that the explosive in a borehole does not detonate well because either the quality of the explosive is low or the charging operation of the explosive is incorrect. For example, some large air gaps are left in the blastholes due to too-fast feeding. 6. Too-large initiation errors for the neighboring holes. For example, the hole planned to initiate earlier is actually initiated after the hole planned to initiate later. This may occur when pyrotechnic detonators are used. 7. Connection between detonator wires and detonating cord. For example, when a double knot is not used to connect a detonator wire and a detonating cord, the detonator may not be initiated, as described in the chapter: Explosives and Detonators. 8. Bad detonators. This seldom happens but it does sometimes occur. 9. Incorrect detonator position. As analyzed in the chapter: Primer Placement, in normal blast condition, it should be avoided that a detonator is placed at or close to the collar of the borehole. 10. Simultaneous initiation of two or more neighboring blastholes. In the Ridgeway mine, two middle holes would be first initiated at the same time in a ring, and then their neighboring holes should follow. But the second delay often had a malfunction [11]. This initiation often results in the malfunction to the neighboring holes since such a simultaneous initiation produces a much larger fragmentation radius Rf than a single hole initiation. 11. Borehole damage due to the variation of in situ stress field. According to borehole filming investigation, the borehole damage includes various forms such as deformed borehole, sheared borehole, broken-out borehole, and jammed borehole [12].
19.10 FRAGMENTATION, ORE RECOVERY, AND MINING PROFIT As mentioned previously, the main disadvantages of sublevel caving are high ore loss and high dilution, and both can be up to 20% or even more in large-scale sublevel caving mines. If an iron ore mining company has an annual ore production of 30 million tons with a grade of 64%, the total ore loss per year will reach 6 million tons of iron ore. What a figure for ore loss! We will see that rock fragmentation can be largely improved by blasting, and a better fragmentation can increase both productivity and ore recovery and finally raise the profit of a mining enterprise. These subjects will be described in the chapter: Effect of Blasting on Engineering Economy.
388 PART | V Rock Blasting in Engineering
19.11 SAFETY, ENVIRONMENT, AND VIBRATION CONTROL 19.11.1 Safety Working safety in underground mining is related to various factors such as blast operation, rock conditions, and mining methods. Here we do not discuss the accidents such as fires, traffic casualties, ventilation, water, electricity problems, and so on. Only blast-related safety problems are accounted for here, and some of them are roof problems, rock spalling, rock bursts, rock falls, and brow damage. Excluding brow damage, which has been discussed in this chapter, the others will be described in the chapter: Safety in Rock Engineering.
19.11.2 Environment It is a common problem that rock blasting in underground mines makes local underground water contaminated more or less. There are mainly two reasons for the problem. (1) In multihole blasting, the explosive in one or more holes is not detonated. The undetonated explosive will partly go into the local underground water system, and partly be transported to mineral processing plants together with ores. (2) When an upward hole is charged, for instance in a sublevel ring, the explosive sinks in boreholes and then falls to the floor during and after charging, as shown in Figs. 19.15 and 19.16, which frequently happens in production charging, especially when emulsion is pumped into a wet hole. The main reasons for explosive sinking are (1) great borehole deformation due to the variation of the local stress field, (2) water in boreholes, and (3) low stickiness of the explosive. Therefore, in order to avoid explosives falling out of b oreholes,
FIGURE 19.15 Precharged explosive sank to collars. (a) Upper parts (over 15 m from collars) of boreholes in sublevel caving ring were precharged; after a few weeks the explosive sank to the collars; (b) one borehole from the ring. The pictures show the roof of the production drift.
FIGURE 19.16 Explosive fell to the drift floor after it was loaded into the blastholes.
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precharge should not be used in the cases where boreholes are either wet or in a large deformed rock mass. In a ddition, to reduce incidents of falling explosives from precharged boreholes, stickiness of an explosive should be increased.
19.11.3 Ground Vibrations The blast-caused ground vibrations will be described in the chapter: Reduction of Ground Vibrations.
19.12 SUBLEVEL-CAVING BLASTING IN THE FUTURE The mining industry has the following missions: l
To increase ore recovery or reduce resources loss in mining operation; To increase profit in the mining process; l To secure safety in underground mining. l
In order to complete these missions, the following are important measures: l
l l l l l l l l l l l
Correct primer placement; Correct stemming; Proper charge diameter to meet the requirement for sufficient VOD; Use of dividing single blasting (DSB) method if necessary; Appropriate control of ore flow such as by leaving an ore layer as a buffer; High specific charge or smaller burden; Double-ring blasting; Efficient open cut and drifting; Correct mining layout and mining sequence; Increase in ore recovery from previously finished sublevels and some single rings where ore extraction is going on; Interfering high-stress state or releasing high stresses in the high-stress regions; Reducing borehole damage.
Some of these measures were described in the previous chapters such as primer placement, but others are not well developed. For example, the mature techniques on interfering high-stress state and reducing borehole damage are few. This is because sublevel caving mining meets with more and more challenges as mining depth increases. Some of the challenges are: l
High-stress state in deep mining. With increasing mining depth, the stress state becomes higher and more severe. The high-stress state can result in different problems, some of which are: (a) large borehole deformation which causes the precharged explosive to be pushed out of the borehole; (b) rock bursts or seismic events, especially when a fault is involved; the rock bursts or seismic events can seriously damage production drifts, including boreholes; (c) various problems of borehole damage [12] (see Fig. 19.17). These problems largely undermine rock blasting and adversely affect fragmentation and ore recovery. l Effect of hanging wall collapse on borehole stability. Each time when a hanging wall suddenly collapses or caves in, a great amount of energy in the form of stress waves is released, partly into the solid rock mass and partly into the previously caved waste rocks. In consequence, the stress field and stress state in the mine will be affected. The rapid change in the stress field and stress state can result in borehole deformation and damage. l Effect of production drilling on borehole stability. Production drilling is often done a long time, for example, one year, before rock blasting. This means that boreholes are loaded a long time before they are charged. Therefore, when this time is shortened, the borehole deformation and damage may decrease. This is to be confirmed by further study. l Effect of blasting on muckpile’s compaction and ore recovery. Fig. 19.18 shows a hanging-up and rock–ore sandwiches in sublevel caving mining. Probably, a simultaneous blasting of two or three middle holes is a major reason for the hanging-up and the compaction. This compaction adversely affects ore recovery, and the hanging-up has a negative effect on production and safety. l Optimum fragmentation from mine to mill. Optimum fragmentation has a great economic potential in underground mining. It may be realized by two ways. One is to increase energy efficiency in each operation such as drilling, blasting, crushing, and grinding. For example, to increase energy efficiency in blasting by means of efficient stress superposition. The other way is to increase the energy input to the blasting operation, so that the energy consumption in downstream operations such as crushing and grinding is dramatically reduced. A detailed description of optimum fragmentation will be presented in the chapter: Optimum Fragmentation.
390 PART | V Rock Blasting in Engineering
FIGURE 19.17 Borehole instability problems in sublevel caving. (a) Production blasthole is sheared, broken, and jammed; (b) production blasthole is fractured.
FIGURE 19.18 Hanging-up and compact ore–rock sandwiches in sublevel caving. The middle part of the picture is the drift. The place at which the lamplight points is a layer of waste rock. Over and below this layer is a layer of ore left. Another layer of waste rock can be seen close to the upper part of the drift.
19.13 CONCLUDING REMARKS 19.13.1 Characteristics of Sublevel Caving Blasting The main disadvantages of sublevel caving are high ore loss and high dilution, and both ore loss and dilution can be up to 20% or even more. However, ore loss can be reduced by improving rock blasting. The spacing varies to a great extent, which results in an uneven distribution of explosive within the ore mass to be blasted. Accordingly, spacing, charge length, and initiation plan are to be determined carefully.
19.13.2 Subdrilling, Deviation, and Inclination As in open pit blasting, subdrilling in sublevel caving is also necessary for both a narrow ore body and the rings directly under hanging walls. This is often overlooked. In the case of a large deviation in rock drilling, delay time and initiation sequence have to be adjusted so as to consider the actual borehole position in the final blast operation. An inclined ring is better for brow protection than a vertical ring.
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19.13.3 Single-Ring Blasting and Double-Ring Blasting For a narrow ore body, it is better to choose single-ring blasting instead of double-ring blasting so as to reduce permanent ore loss. For a large ore body, double-ring blasting with an effective extraction control should be favorable for fragmentation and ore recovery.
19.13.4 Delay Time The principles for determining delay time described in the chapter: Delay Times are valid for sublevel caving blasting. In order to obtain optimum delay time, either detonation wave (or borehole pressure) or stress wave close to the blasthole is necessary. At the same time, electronic detonators are recommended.
19.13.5 Brow Damage Serious brow damage not only gives rise to lower ore recovery and higher dilution, but also undermines safety for the people working in the field. Brow damage can be reduced by taking various measures such as multiprimer placement in a blasthole, correct detonator position, reduction of specific charge, high-VOD explosive, proper initiation plan, and so on.
19.13.6 Misfire Misfire is still a serious problem in sublevel caving and other types of underground mining. The main causes of misfire are pressure transients, too-small spacing, sympathetic detonation due to fractures or empty rooms within the rock mass between two holes, too-long delay time between neighboring holes, low quality of explosives, incorrect charge operation, wrong detonator position, and so on.
19.13.7 Other Issues Related to Sublevel Caving Blasting There are other important issues related to sublevel caving blasting such as open cut, drifting, stemming, air deck, detonator placement, fragmentation, mining cost, mining safety, and environmental influence. Some of these have been described previously, and the others will be discussed later in the book.
19.13.8 Underground Blasting in the Future Underground blasting in the future will have to pay attention to problems caused by a high-stress state, such as borehole instability, high confining pressure in blasting, effect of blasting on muckpile and ore flow, and optimum fragmentation.
19.14 EXERCISES 1. List the advantages and disadvantages of the sublevel caving method. If you are assigned to a project aiming to reduce the ore loss in an underground mine where sublevel caving is used, do you know which methods can be used to reduce the ore loss in sublevel caving? 2. In Fig. 19.1, if blasthole 5 is not detonated, which consequences can be caused for mining production and environmental protection? 3. In Fig. 19.2 the subdrilling is necessary. If there is no subdrilling in the rings under the hanging wall, what consequence may be caused for the mining production? 4. In the cases of deviation shown in Fig. 19.4c and d, how do you make your blast plan in order to ensure that all of the holes are detonated? 5. In which case may a double-ring blast be better for fragmentation and ore recovery? To ensure that the recovery is high in a double-ring blast, what should you do? 6. List the measures for reducing brow damage in sublevel caving. Explain each measure by using your own words. 7. Use the stress wave theory in the chapter: Stress Waves to explain why there is much spalling in the roofs shown in Fig. 19.14a. 8. In normal conditions, the emulsion explosive pumped into a blasthole can well stay in the hole for a long time (up to two months) without sinking. But sometimes or in some mining areas, the explosive charge markedly sinks, as shown in Fig. 19.15. Give the possible reasons for the sinking. 9. Use Eqs. 19.1 and 19.2 to prove this relation: Bebv > Bebi.
392 PART | V Rock Blasting in Engineering
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