SME Mining Engineering Handbook 2nd Edition Volume 2
Senior Editor Howard L. Hartman Professor Emeritus of Mining Engineering The University of Alabama
Associate Editors
Scatt G. Britton
Jan M. Mutmansky
Vice President Tanoma Mining Co.
Professor, Dept. of Mineral Engineering The Pennsylvania State University
Danald W. Gentry
w. Joseph Schlitt
Head, Dept. of Mining Engineering Colorado School of Mines
Manager of Technology Minerals, Metals, and Chemicals Brown & Root Braun
Michael Karmis Professor and Head, Mining Engineering Virginia Polytechnic Institute and State University
Madan M. Si'ngh President Engineers International, Inc.
Cosponsored by Seeley W. Mudd Memorial Fund of AIME Published by Society for Mining, Metallurgy, and Exploration, Inc. LitUeton, Colorado • 1992
Chapter 22. 1 RAPID EXCAVATION C.D. BREEDS AND
J.J.
CONWAY
in this chapter is, therefore, to present a list of the main components importan-t to estimating project costs and to direct the reader to potential unit cost providers.
22.1.1 INTRODUCTION For the purposes of this section on innovative mining methQds, rapid excavation is defined as underground excavation by means faster than conventional methods. However, many of the techniques and mining methods described below are well proven in civil construction and in a small nUITlber of US mines. A major innovation would be broader acceptance of these technologies by the mining industry. Two major organizations in the United States promote the use of rapid excavation techniques for civil and mining applications. The Executive Board for Rapid Excavation and Tunneling Conferences (RETC) was established in 1971 to disseminate technical information in this rapidly advancing field of underground construction. The RETC and its proceedings provide a wealth of case study information related to site investigation, groundwater control, design and analysis, equipment, instrumentation, materials handling, and support for rapid excavation projects in soft ground and hard rock. The more recently established Institute of Shaft Drilling Technology (ISDT) provides a forum for discussing and reporting advances in shaft drilling. Short courses in mining techniques, shaft sinking, and boring techniques are provided through the ISDT and are highly recommended for engineers and owners planning major shaft projects. This chapter draws extensively from publications of these organizations, field experience in rock cutting and excavation engineering, and input from equipment manufacturers and contractors. Each segment has been written to provide the reader with a description of the equipment used and an overall appreciation of selection methodology. Emphasis is placed on methods and equipment used for mine access construction and mine development. Rapid excavation methods associated with development and production mining (e.g., longwall mining, continuous mining, and stoping methods) are discussed elsewhere in the Handbook (see Chapters 17.4, 17.5, 18.1, 18.2, 19.1, and 20.1).
22.1.2 MECHANICAL ROCK CUTTING TECHNIQUES AND THEIR APPLICATION TO MECHANICAL MINING EQUIPMENT The mechanics of mechanical rock breakage, and the parameters important to determining cuttability and production rates are presented in Chapter 9.2 of this Handbook. The objective of this chapter is to describe five basic cutting methods and their application to mechanical mining equipment. These basic cutting methods, defined in terms of tool type, are illustrated in Fig. 22.1.1 and include: 1. Drag bit cutting. 2. Point-attack bit cutting. 3. Disk cutting. 4. Button cutting. 5. Roller cutting.
22.1.2.1 Drag Bit and Point-attack Bit Cutting The application of both drag bits and point-attack bits is similar. The tools are inserted in tool holders (or boxes), which are· integral parts of the cutting head, and may be held in place by a circlip or spring. Point-attack bits are commonly free to rotate in their holders. It has been claimed that this feature promotes more even tool wear (self sharpening) and better overall tool life, although research by Hurt and Evans (1981) disputes this. During cutting, the bits are pushed into the rock, developing cutting forces parallel to the direction of head rotation and normal forces parallel to the direction of head thrust. As these forces build up to critical values, a macroscopic failure surface develops ahead of the bit, and a piece of rock spalls away. The pick then moves ahead into the space left by the spalled chip until a new rock buttress is encountered, and tool forces again build up. The cutting process is thus a cyclical one with rapid fluctuations in tool forces. Adjacent bits produce parallel grooves and interaction between these has an important influence on cutting efficiency. Roadheaders use drag and point-attack bits almost exclu~ively. These tools also find application on tunnel boring machine (TBM) cutterheads, but in this role they are generally limited to machines operating in weaker formations.
22.1.1.1 Rapid Excavation System Performance A short section on system performance evaluation is provided for each rapid excavation method described. Simple empirical techniques, which utilize existing case study data and qualitative information, are used to estimate the probable range of system performance. This approach is considered to be applicable at a conceptual level of project planning. More detailed analyses, rock cutting experimentation, and equipment/system performance predictions are available from equipment manufacturers, but, due to space constraints cannot be adequately dealt with here.
22.1.2.2 Disk Cutting 22.1.1.2 Cost Estimating
Disk cutters (Fig. 22.1.1 c) generally consist of solid steel alloy discs with a tapered cutting edge. The disk is mounted in a bearing and is free to roll in response to applied forces acting parallel to the rock surface. These rolling forces are analogous to the cutting forces applied in drag bit cutting. Thrust and drag forces are applied to the disk through the bearing and act normal and parallel respectively to the rock surface. Disks used in practice may be of the simple type illus-
Since the inception of mechanized mining, many papers have been published which enumerate the absolute cost advantage of mechanical vs. conventional construction. However, technical advancement in equipment design, owner experience, and increasing competition among contractors decreases the utility of absolute cost estimates especially when presented in a medium with an anticipated useful life of a decade or more. The approach
1871
1872
MINING ENGINEERING HANDBOOK
A. Drag Cutter
B. Point Attack Cutter
C. Disk Cutter
D. Roller Cutter
E. Button Cutter
T-Thrust FN - Normal Force Fc - Cutting Force F R- Rolling Force Fig. 22.1.1. Rock cutting techniques (after Roxborough and Rispin, 1973).
RAPID EXCAVATION trated, or may consist of multi-edge varieties, including types with successively smaller disk diameters giving a tapered or conical arrangement. Frequently these multi-row disks employ carbide inserts with chisel points imbedded nearly flush with the circumference. Simple disk cutters are used primarily on full face TBMs, and multi-row disks on raise boring machines (RBMs). Thrust forces acting on the cutting head push the cutter into the rock building up stresses which cause local rock failure. Because of the translatory motion of the cutting head, the disk rolls forward cutting a groove in the rock. As in the case of drag cutters, ~interaction between adjacent grooves has been shown to have an important influence on cutting efficiency.
22.1.2.3 Roller or Mill Tooth Cutting Roller or mill-tooth cutting is similar to disk cutting except that instead of a tapered disc edge, the tool is equipped with circumferential teeth (Fig. 22.1.1 d). As the cutter moves in response to rolling forces, each tooth in turn is pushed into the rock, acting like a wedge, and causing local failure.
22.1.2.4 Button Cutting Button cutters consist of cylindrical or conical tool bodies inset with tungsten carbide buttons (Fig. 22.1.1 e). The tool is mounted in a bearing in the same way as disk cutters or roller cutters and is free to roll in response to applied forces acting parallel to the rock surface. Thrust forces cause high stress concentrations beneath each button as they roll across the rock surface, resulting in local failure and pulverization of the rock. The area of influence of each button is small and results in a fine-grained product. Because the product size is small, specific energy requirements are high and button cutting is the least efficient of the rock cutting methods discussed. Button cutting is used in applications in which high rock strength and abrasivity preclude the use of other methods. These cutters also find application as reaming cutters used for final profiling on RBMs and TBMs.
22.1.3 BASIC METHODS OF PREDICTING INSTANTANEOUS CUTTING RATES 22.1.3.1 Introduction When considering the feasibility or cost effectiveness of employing a mechanical excavation system, the central questions are (1) Can this machine cut this rock? (2) If so, how fast? and (3) What is the cost of maintaining this performance? Clearly there is a need for a reliable method of performance prediction. Two aspects of machine performance need to be assessed to answer the above questions. First, machine performance in terms of cutting rates or penetration rates must be assessed. Second, the overall system performance and reliability, with particular reference to those aspects that impact machine utilization, must be assessed. In the following discussion, methods of predicting or estimating cutting rates or penetration rates will be described, while methods of overall system assessment will be addressed in subsequent segments dealing with specific mechanical excavation methods. Prediction of cutting rates requires information on rock material properties, rock mass properties, and machine characteristics. The link between these three groups of data is provided by what may be termed rock-tool or rock-machine interaction models, and the result of applying such a model is an estimate
1873
or prediction of performance. In the following discussion, prediction methods are placed into two broad categories depending on whether the interaction model is theoretical or empirical. Before discussing performance prediction, the following terms must be defined: Cutting rate (used in conjunction with roadheaders and boom-type tunneling machines) is the rate at which rock is excavated during cutting (volume excavated/cutting time), usually expressed in units of ft 3/hr (m 3/h). Care must always be taken to determine whether quoted "cutting rates" refer to what may be termed the instantaneous cutting rate (ICR) or the operational cutting rate (OCR). Cutting rates determined under highly controlled conditions, such as a research field test, in which cutting time is recorded as the actual time spent in cutting (determined from instrument measurements of power consumption against time) are instantaneous cutting rates. Under typical operational conditions, cutting time is generally taken as synonymous with utilization. Minor delays resulting, for example, from adjusting the boom position at the end of each cutting traverse, or reduced rates of production during final profiling, are neglected. Cutting rates determined using utilization as the cutting time are termed operational cutting rates. Clearly, performance predictions based on instantaneous cutting rates, without an appropriate cutting time correction, will be overly optimistic. Back analyses suggest that operational cutting rates commonly have values in the range of 0.45 to 0.60 times the instantaneous cutting rate. For final profiling, this figure may drop to 0.3, while during bulk production, an experienced operator may achieve a ratio as high as 0.85. Specific energy is a commonly used measure of cuttability that is defined as the work done to excavate a unit volume of rock. In the context of rock cutting, specific energy should not be thought of as a fundamental property of the rock. Rather, it is a function of rock properties, cutting tool design, and cutting tool interaction, in the same way· as compressive strength is a function of specimen size, shape, and test conditions. Measured specific energies are many times greater than theoretically determined values, the difference being accounted for in energy lost to frictional heating, vibration, and so on. Penetration rate (used in conjunction with full-face shaft or tunnel boring machines) is the rate of advance measured during the cutting cycle, normally expressed in inches or feet (meters)/ revolution or feet (meters)/hour. For practical purposes, instantaneous and operational penetration rates are considered equal. Utilization is the time remaining for excavation when planned and unplanned machine stoppages have been accounted for. Stoppages are required for a variety of reasons including support installation, survey work, pick replacement, routine and non-routine maintenance, .muck haulage delays, shift changes, and so on. Advance rate is the rate of tunnel or drift advance, usually expressed in units of feet (meters)/day, feet (meters)/shift, etc., and is equal to OCR/face area X utilization or penetration rate X utilization (22.1.1)
22.1.3.2 Theoretical Models of Rock Cutting Theoretical models have been proposed that attempt to analyze peak forces required, or work done, to excavate a unit volume of rock, and to relate these to fundamental rock properties such as shear and tensile strengths and internal friction angles. All these models have certain weaknesses that limit their usefulness for solving practical problems in machine design and performance. These weaknesses relate to a poor understanding
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MINING ENGINEERING HANDBOOK
of both the state of stress developed in the rock as a result of the applied forces and the mechanics of crack initiation and propagation. In addition, materials are generally considered to be homogeneous, and the important influence of pre-existing fractures is ignored. Even with the simple case of a single cutting tool, a complex three-dimensional state of stress must exist in the rock around the tool tip. It is generally acknowledged that in the immediate contact area of the tool, intense crushing of the rock must occur, and that the properties of this crushed material differ markedly from those of the "intact" material. Theoretical approaches generally assume a simplified two-dimensional stress distribution, such as a point or line load, and neglect the properties of the crushed zone and the important role of this zone in transmitting stresses from the tool to the intact rock. Further, in practical cutting applications, multiple tools are arranged in a manner that promotes interaction between adjacent cuts, which has been shown to improve the overall efficiency of the system. This introduces a further level of complexity to the three-dimensional stress distribution that tends to be neglected in theoretical models. Both brittle and plastic failure modes have been considered in theoretical rock cutting models, the appropriateness of each depending on the initial properties of the rock, and changes induced during cutting. Even in brittle rock, plastic deformation may occur in the intensely stressed zone adjacent to the tool tip. Failure criteria based on both tensile and shear stresses has been applied to rock cutting, although in practice, failure may be initiated in one mode and change to the other as the stress distribution changes during crack propagation. Thus a rigorous theoretical description of rock cutting must incorporate a sophisticated failure model, which accounts for both localized differences in material behavior and transient responses to a changing stress distribution. However, there can be little justification for developing or applying such a failure model until equally sophisticated three-dimensional stress distribution models are available. ApPLICATION OF THEORETICAL CUTTING MODELS TO ROADHEADERS. In the case of roadheaders, the limitations of theoretical models are compounded by the relatively large number of pick geometries available, the mode of roadheader operation (which involves continually varying normal forces), depths of cut and mode of cutting [Le., sumping, traversing, etc. (Fowell and McFeat-Smith, 1976)], and a generally less-controlled cutting environment. Cutting theories applicable to roadheaders are not considered sufficiently developed at this time to be useful as prediction tools and are not discussed further here. ApPLICATION OF THEORETICAL CUTTING MODELS TO TUNNEL AND SHAFT BORING SYSTEMS. In the case of full-face excavation systems, theoretical modeling problems are less acute. Here, variations in cutter geometries are limited to variations in disk diameter and blade width. In addition, the cutting process is more controlled, involving relatively constant penetration rate and depth of cut, and only a single cutting mode. Because of this, some progress has been achieved in the application of theoretical cutting models, albeit oversimplified, to prediction of the performance of full-face TBMs. The better theoretical models of TBM performance are widely used as prediction tools, however, occasionally a significant deviation occurs. Whether the problem is in the model or in the ability of the sample or geotechnical data to represent the rock mass is not clear. To be useful, such models must be able to predict thrust forces and rolling forces corresponding to specific depths of penetration in relieved cutting. Conversely, the models may predict achievable penetration given machine constraints governing available thrust and rolling forces. A model of this type will
predict machine advance per revolution for a given machine power and tool spacing; a separate calculation of yield per revolution is not required. Roxborough and Phillips (1975) have presented expressions for thrust force Ft and rolling force F" acting on a disk during unrelieved cutting:
Ft = 4o-c X tan 8/2 X (Dp 3 - p4)O.S Fr
= 4o-c
X p2 X tan 8/2
(22.1.2) (22.1.3)
where o-c is unconfined compressive strength (UCS), 8 is disk edge angle, D is disk diameter, and p is depth of penetration. Based on breakage patterns observed during actual cutting tests, they concluded that the failure process is controlled by shear stresses acting on the plane connecting the apices of adjacent grooves. Comparison of experimentally determined forces (for Bunter sandstone) with calculated values presented by these workers indicated good correlation. Farmer and Glossop (1980) have presented these equations (slightly modified in the case of Ft), and claim that expressions of this general form are in reasonable agreement with experimentally determIned results. Roxborough and Phillips (1975) also suggest that the optimum spacing/penetration ratio is given by (22.1.4) where T is shear strength of the rock. Again, good correlation was demonstrated between calculated and observed Sip ratio for Bunter sandstone. Eqs. 22.1.2 to 22.1.4, however, provide only a partial solution for prediction or head design. Using Eq. 22.1.4, the optimum spacing for a given penetration p can be calculated. Using this value of p, it should then be possible to calculate Fr and Ft for individual tools, using Eqs. 22.1.2 and 22.1.3. The total number of tools can be determined from the optimum spacing and head diameter, and hence the total torque and thrust requirements can be determined. However, these will be overestimated because Eqs. 22.1.2 and 22.1.3 apply to unrelieved cutting, whereas the actual spacing is selected to minimize tool forces. Because of the current limitation of theoretical models, practical design approaches use empirical methods, as described in 22.1.3.3.
22.1.3.3 Empirical Methods of Predicting Instantaneous Cutting Rates for Roadheader and Boom-type Tunneling Machines Because of the theoretical difficulties of modeling roadheader cutting performance, approaches to this problem are essentially empirical. It can be claimed that theoretical considerations have shed some light on which material and machine parameters have an important influence on performance, but while these parameters appear in many empirical performance equations, they are always associated with dimensionless constants derived from actual cutting trials or performance data. The simplest empirical prediction methods are based on the extrapolation of performance records of specific roadheader models under specific geotechnical conditions that match those of the proposed site. While this approach has the ~dvantage of simplicity, it also has a number of weaknesses. It is very difficult to collect high-quality roadheader performance data under other than the highly controlled conditions of a research project. Performance data collected under typical operational conditions,
RAPID EXCAVATION
1875
45 40 o
35
Observed
- S E = 115/CR
M
.E 30 ...., ~
>.
~ Cl>
c::
UJ
20
~ ·0
15
Cl>
C. Cl)
10 5 o
40
60
80
100
120
140
Instantaneous Cutting Rate (m 3 /hr)
Fig. 22.1.3. Comparison of instantaneous cutting rates and specific energy requirements for a DOSCO MKIIA (after McFeat-Smith and Fowell, 1977). Conversion factors: 1 Btu/ft3 = 0.0373 MJ/m3, 1 ft3/hr = 0.0283 m3/h. o
10
15
20
In situ specific energy MJ/m3 18.....---------"---------------------,
,....
Fig.22.1.2. Correlation of laboratory specific energy and in situ speci'fic energy (after McFeat-Smith and Fowell, 1977). Conversion factor: 1 Btu/ft3 = 0.0373 MJ/m3.
ROD
16
,
"
0-30,0,%,./
14
SE =HP/ICR
(22.1.5)
where SE is specific energy, HP is head power, and ICR is instantaneous cutting rate. This curve provides a very good upper bound fit to the measured data, and in most cases shows that actual specific energy was less than predicted to achieve a given cutting rate. This may reflect the tendency of rock mass structural features to reduce specific energy requirements. Fig. 22.1.4 shows a plot of observed OCRs from various sources, vs. predicted cutting rates using McFeat-Smith and
8
l
~
'
10
2530
8
0
'0
0
22
0,
6 4
~ ~ •• .,.,.,.,.,.,,,
ROD c 30 - 100%
a: '12
which constitutes the bulk of the data base, must be treated with caution. Furthermore, a good match between geotechnical conditions at the proposed site and a past site may not exist. In this case, the process becomes rather subjective, and there is no clear means of deciding what weight to attach to particular parameters. In an effort to remove some of the subjectivity and identify important performance predictors, McFeat-Smith and Fowell (1977) investigated the relationships between rock index properties, laboratory specific energies (determined from small-scale cutting tests), in situ specific energies (determined from fieldscale cutting tests),. and instantaneous cutting rates for a variety of British Coal Measure rocks. Application ofmultivariate statistical methods to the results of laboratory tests enabled these workers to derive prediction equations that use a small number of index properties to predict specific energy requirements for rock cutting. These predictions were shown to correlate well with field specific energy measured during actual cutting trials (Fig. 22.1.2). Field specific energy was shown to be related to cutting rate using a very simple rock/machine interaction model. Fig. 22.1.3 shows a plot of measured in situ specific energy against cutting rates for Coal Measure strata reported by McFeat-Smith and Fowell (1977). Included on this plot is a theoretical curve developed from the rock/machine interaction model:
~~
75
21
'70
,"....Jo
•
100
100
~~
75 • .,.,.,.,.".,
50
5~ +
60
2~' '
30 ,
~~~100
,,"
..
:r' s:!
~
~~~
"
100 ....100
50 60+ ~
~~
9 . . 6~ ~~~O
2 ~+~~~ , ,.~60
O~__IpL___._____r"-..__....--....._____._____r"-..___r_~____r_____r-..____......__~____r____I
o
2
4
6
8
10
12
14
16
18
Predicted OCR
Fig. 22.1.4. Comparison of observed operational cutting rates and operational cutting rate predicted using data from McFeat-Smith and Fowell.
Fowell's predictive equations and the simple rock/machine interaction model given by Eq. 22.1.5. Once again, in nearly all cases, actual performance was better than predicted. When rock quality designation (RQD) is considered, the data are seen to fall into two broad fields, although considerable scatter is still present. However, those data points for 100% RQD fall close to the lower bound (i.e., predicted = observed). Many of the points included in Fig. 22.1.4 are for roadheaders with up to twice the cutting-head power of the machine utilized in McFeat-Smith and Fowell's work, and cutting rates for this machine were predicted simply by inserting an appropriate value of HP in Eq. 22.1.5. It would appear, therefore, that these predictive equations may be applicable to a range of machines, provided that appropriate cutting time factor corrections are made. Also, direct determinations of specific energy using core grooving tests could be used in conjunction with Eq. 22.1.5 to predict instantaneous cutting rates. The predictive equation approach has also been used by Aleman (1983), who has demonstrated good correlations be-
MINING ENGINEERING HANDBOOK
1876
.s/
3-
2-
," • ·C • --r.. -.~- /
1-
A'
/
/
/
100
/
~
/. /.
O~---,-----,----...,....----...,....-----.-----r--'---'
o
~~ O~---~I-----rl------'-I---~
o
123
4
Actual performance nf/kN
2
3
6
5
4
Bits/Foot (e) and Feet/Hour of Machine Cutting p)
Fig. 22.1.6. Roadheader performance vs. rock class, P.21 A Test, 2375 Level (after Sandbak, 1985). Conversion factor: 1 ft = 0.3048 m.
Fig.22.1.5. Comparison of observed and predicted roadheader performance using Aleman's method (after Aleman, 1983). Conversion factor: 1 ft3/lbf = 6.6367 m3/kN. 20
tween predicted and observed performance for a variety of roadheaders (Fig. 22.1.5). Important aspects of Aleman's approach are the inclusion of RQD and an assessment of microfracturing in the predictive equations so that a secondary assessment of the influence of rock mass condition is not required. The predictive equations are used in conjunction with a more sophisticated machine model than that of Fowell and McFeat-Smith. This approach takes into account the limitations of available arcing force, and head rotation speeds, to derive instantaneous cutting rates. It is worth noting at this point that many of the indices or parameters that appear in predictive equations, or are referred to in the literature as being significant predictors of roadheader performance, are often strongly correlated. Several parameters show strong correlation with unconfined compressive strength. Therefore it is not surprising that this parameter is shown to be significant in most of the studies undertaken. Where such correlations can be demonstrated, some predictive equations can be simplified to give expressions primarily in terms of unconfined compressive strength. With good data collection, cutting rates can generally be correlated with rock mass and rock material properties at specific sites. The results of this type of study provide a useful means of predicting performance of a specific machine type under a variety of geological conditions. But since the results are not presented in terms of specific energy (McFeat-Smith and Fowell, 1977) or machine characteristics (Aleman, 1983), the results cannot, strictly, be extrapolated to machine types other than those for which they were derived. Two good examples of this type of study have been reported by Sandbak (1985) and Bilgin et al. (1988). Sandbak demonstrated correlations between performance (operational cutting rate and bit consumption) and Bieniawski's rock mass rating (RMR) for a Dosco SL-120. Although the scatter in the results is rather large (Fig. 22.1.6), the overall trends are clear. Cutting rates are lowest in strong rock with few fractures, corresponding to high RMR values, and highest in
16
€
('l)
E a)
co Cl:
12
x xx
0)
c:
'= <3
8
4
O-\--~--r--.,...--r----,--...,-------,--,-...,....----,----,---~----j
o
200
400
600
800 2
Rock Mass Cuttability Index kglcm (ac •
1000 RQ0
1200
2l3
100 )
Fig. 22.1.7. Relationship between machine advance rate and rock mass cuttability index (after Bilgin et al., 1988). Conversion factors: 1 ft3/hr = 0.02832 m3/h, 1 psi = 0.0703 kg/cm 2 •
low-strength, heavily fractured rocks corresponding to low RMR values. Bilgin et al. (1988) collected detailed data on machine performance (for a Herrenknecht SMl), rock mass properties and rock material properties. Statistical analyses showed significant correlations between operational cutting rate and the product of UCS X RQD2/3/100 (or rock-mass cuttability index RMCI, Fig. 22.1.7). These workers also investigated the applicability of the RMCI to prediction of cutting rates for a Dosco Mk 2A and a Pk 2r at other sites and found a reasonable correlation. These results suggest that the RMCI may be applicable to a variety of machines provided that an adequate allowance can be made for variations in head power. It should also be noted that the
1877
RAPID EXCAVATION Table 22.1.1. Summary of Empirical Roadheader Performance Prediction Methods Reference Fowell and McFeatSmith 1976, McFeatSmith and Fowell, 1977
Rock Material Properties
Rock Mass Properties
Machine Characteristics
Rock-Machine Interaction Model
Comments of Applicability
(1) Cone indentor hardness, shore hardness, Cementation coefficient, UCS (2) Cone Indenter hardness, 'plastic hardness.' Unconfined compressive strength (UCS)
Not considered
Head power
Empirical-predicts specific energy and instantaneous cutting rate.
(i) Prediction equations can be expressed largely in terms of UCS. (ii) Needs CTF correction for OCR. (iii) Developed for DOSCO MK IIA, may be applicable to other light-medium-duty roadheaders.
Rock quality designation (RQD)
Not considered
Sandbak, 1985
UCS
Rock mass rating (RMR)
Not considered
Developed for shieldmounted Herrenkneckt SMI. Needs correction for other machine powers and non-shielded or non-stelled machines. peveloped for DOSCO SL 120. Would need correction for other machine powers/head configurations.
Aleman, 1983
UCS Cerchar abrasivity
RQD microfracturing
Torque, arcing force, head RPM
Farmer and Garritty, 1987
UCS
Deformation modulus
Hurt et al., 1981
Not considered
Not considered
Head power, Energy transfer ratio. Number & orientation of tools, radial distance, torque, and arcing forces.
Empirical and machine specific. Predicts OCR for lightweightmediumweight road headers. Empirical and machine specific. Predicts OCR for lightweightmediumweight roadheaders. Empirical-uses machine specific constants linked to geotechnical properties in prediction equations for ICR. Empirical-predicts cutting rate for selected energy transfer ratio. Predicts ICR if tool forces are known.
Bilgin et al., 1988
Herrenknecht machine was shield mounted, a condition which generally results in a more rigid system and hence higher cutting rates than for a non-shielded machine. The methods discussed above tend to be machine specific and/or utilize empirical constants that are machine specific. Other approaches to performance prediction, which are based on fundamental material properties, have been proposed. Farmer and Garritty (1987) make use of a strain energy approach to predict roadheader performance. Input data for this model includes DCS and deformation modulus of the rock mass. The machine model consists of head power coupled with an energy transfer ratio, which accounts for energy loss as a result of frictional heating, vibration and so on. Results indicate that as little as 1 to 2% of the available energy is actually used in rock breakage. Table 22.1.1 summarizes the main features of the prediction methods discussed above, including limitations on applicability. Currently, the best procedure for roadheader performance prediction may be the use of one or more of the empirical methods discussed above, provided that the limitations of these methods are understood. All of the approaches outlined above are
Derivations of empirical constants- not presented in cited references.
Requires detailed knowledge of head design, machine characteristics.
deterministic in nature; however, actual case history data indicate considerable scatter even within a single rock type, which may approximate to normal or logri.ormal distributions. This is because of the inherent variability in rock material and rock mass properties and in operational conditions (including machine and pick condition, operator skill, etc.). A better approach to both the analysis of case history data and the prediction of machine performance is the use of probabilistic methods. Using this approach, geotechnical and operational variables are input to the analysis in the form of probability distributions, the output also being in the form of a probability distribution. EMPIRICAL METHODS OF PREDICTING INSTANTANEOUS CUTTING RATES FOR TBM AND SHAFT BORING SYSTEMS. Reliable empirical approaches to estimating penetration rates for full-face boring machines are generally far less complex than those previously described for roadheaders. Howarth (1986, 1987) reviewed seven published methods for the prediction of TBM advance rates and concluded that one simplified method (Farmer and Glossop, 1980), based on thrust per cutter and tensile rock strength, provided good correlation with actual penetration rates from 20 case histories. A complex model suggested
1878
MINING ENGINEERING HANDBOOK
by Lislerud et al. (1983) was considered to have the potential for more accurate penetration prediction but was discounted on the basis of cost. A brief review of these two methods is provided below. METHOD DESCRIBED BY FARMER AND GLOSSOP (1980). Farmer and Glossop derived a relationship between the average cutter force Fv the penetration per revolution P, and rock tensile strength erif by equating the energy input per unit length of cut to the energy required to satisfy fracture surfaces in the rock: (22.1.6) The value of K was obtained by least squares regression of eight cases where P, Fv and er if were measured, so that in SI units:
mm/rev 8
210 200
1.15 1.10
7
6
190
1.05 Kd
180
1.00 0.95 " 12
170 14"
15-1/2
17"
._LJ
C
.Q
~
160 5
150
Q;
140 130 120 110 100 90 80
c
Q) ~ (,)
4
·in
co
co
3
(22.1.7a) 2
with P measured in mm/rev, FL in kN, and er if in kPa. The equivalent equation using English units can be written:
P = 0.0158 X FL
1----
20
I
30
I
I
I
I
I
40
50
60
70
80
90
Drilling Rate Index, DRI
(22.1.7b)
with P measured in in./rev, FL in lbf and er if in psi. A similar equation has been suggested by Graham (1976) based on use of the Robbins TBM in hard rock with unconfined compressive strengths ranging from 20,000 to 29,000 psi (140 to 200 MPa). 3940FL P=--
kN
Fig. 22.1.8. Basic penetration as a function of drilling rate index (DRI), average thrust per disk, and cutter diameter (after Lislerud, 1988). Conversion factors: 1 in. = 25.4 mm, 1 Ibf = 4.4482 N.
(22.1.8a)
ere!
where, ere! is uniaxial compressive strength in kPa. The equivalent equation written in English units becomes,
Fissure Class
Spacing
5cm
IV
0.1 FL P=--
(22.1.8b)
4
As simple ratios are typically used to relate tensile and uniaxial compressive strength (e.g., ranging from 1 : 10 to 1 : 20), either equation, possibly supplemented by additional performance data, may be used. METHOD DESCRIBED BY LISLERUD (1983, 1988). Lislerud has developed a TBM performance prediction method based on rock mass factors (rock mass jointing, intact rock strength, brittleness, and abrasivity) and machine factors (thrust per cutter, cutter edge bluntness, cutter spacing, cutter diameter, torque capacity and RPM, and cutterhead curvature and diameter). Lislerud's equation for net penetration is written (in SI units):
3
ere!
(22.1.9) where ib is the basic penetration rate in mm/rev and is a function of the thrust per disk and the drilling rate index (DRI) as shown in Fig. 22.1.8 (DRI is based on testing described by the Norwegian Institute of Technology), K d is a correction factor for cutter diameter; and K s is a correction factor for joint rating and frequency (see Fig. 22.1.9). As noted by Howarth (1986), this method requires a considerable amount of geotechnical and laboratory test data and is probably only suited to foliated, high-grade metamorphic rocks such as those found in Scandinavia. In less anisotopic rocks, use of the simpler relationships suggested by Farmer and Glossop (1980) and Graham (1976) is warranted.
Ill-IV
2 III
10cm
II - III I1
J
-0.36
o
10 20 30 40 50 60 70 80 90
at 190 kN
20cm 40cm
a
Angle between tunnel axis and planes of weakness
Fig. 22.1.9. Correction factor Ks as a function of fissure class and angle between tunnel axis and planes of weakness (after Lislerud, 1988). Conversion factors: 1 in. = 2.54 cm, 1 Ibf = 4.4482 N.
22.1.3.4 Summary Empirical methods currently provide the best means of estimating machine performance. Such estimates can be made directly based on previous experience in similar ground conditions or can utilize one of the predictive equations. In considering the results, the user should be aware of the limitations of each method. Input should also be sought from machine suppliers or
1879
RAPID EXCAVATION Table 22.1.2. Case Study Cycle Times for Blind Shaft Drilling D. Runge and J.T. Zeni (1987)
Cased Study Reference Source
H.E. Hunter (1982)
H.E. Hunter (1982)
H.E. Hunter (1982)
Crookston et al. (1983)
Mobilization
(days)
5
12
7
4
Drilled depth Drilled diameter Drilling time
(ft) (in.) (days)
1050 87 & 66 53
2243 120 129
2188 72 64
2188 72 66
2371 120 238
Casing length Casing diameter Prepare for casing Run casing Grout casing
(ft) (in.) (days) (days) (days)
394 67 (inc) 6 (inc)
2194 85 5 21 10
2132 36 2 8 5
2131 36 1 7 7
2371 96
Pump casing
(days)
6
4
2
Maintenance
(days)
7
(inc)
(inc)
(inc)
Total cycle -Drilling -Casing and grout
(days) (ft/day) (ft/day)
71 20 66
183 18 61
90 34 142
90 32 142
Formation
Weak shale
Circulating
Reverse circulation
Sandstone and shales Reverse circulation
Sandstone and shales Reverse circulation
22
Sandstone and shales Reverse circulation
Shale Reverse circulation
Conversion factors: 1 in. = 25.4 mm, 1 ft = 0.3048 m.
. . . : : > : ! : : : " - - - - - - Crown block
specialist consultants before decisions on a machine's suitability for a particular application are made.
D.rrick - - - - - - - 1 1
22.1.4 SHAFT CONSTRUCTION SYSTEMS AND EQUIPMENT
Traveling block Gooseneck
Three major rapid shaft excavation systems are described in this subsection, namely, blind shaft drilling, vertical tunneling mole or V-mole, and raise drilling. Data are included to provide the reader with a means of comparing the relative merits of each system and to assist with method selection.
-----ti-'~
------,~"'""'J_____\_\1r_----
~r--m::::::;;=;:::::== ";
Rotary hose
Rotary drive bbuu~Sh~in~g-=~$3~~~~~~~ Rotary tabl. -
Bail Swivel Drawworks Electrical or engine drive
Substructure K.II Ground l.v.I---....-....."L".,"'lI
e.-+------
22.1.4.1 Rotary Blind Drilling Systems Large-diameter shaft drilling systems are an extension of conventional rotary drilling techniques used extensively for oil well boring. Extensive development work was pioneered by the Atomic Energy Commission (AEC) during the 19608 as part of the US nuclear testing program at the Nevada Test Site. Blind drilled shafts have since provided rapid access for underground mining projects throughout the world and are proven under a wide range of operational and site conditions (see Table 22.1.2 for summarized case study data). System components and operational considerations are described below. A generalized blind shaft drilling equipment set-up is shown in Fig. 22.1.10. SHAFT COLLAR AND FOUNDATION. A shaft collar is typically excavated using either an auger rig or conventional drilland-blast mining during mobilization of the blind shaft drilling equipment. Collar depth depends on the overall length of the bottom-hole drilling assembly and is designed so that the assembly can be positioned below the drilling rig's rotary table. The collar may be lined with steel, shotcrete, or concrete, depending on ground conditions. A cast-in-place, reinforced concrete foundation will be designed to support the drill rig. DRILLING RIG. Major components of the drilling rig include a mast and substructure, drawworks and tugger hoists, rotary table, crown and traveling blocks, hook, swivel and kelly. The
I+\'f-----===- Drilling lin.
Drill pip. tool joint
Drill p i p < r - - - - - - - i - U ..d::j~------
f
Hold down clamp
Upper roller type stabilizer
~
t::::::::::!--If-------==_ Oonut type steel weights
l;; Wall 01 hole -~---o'fi
L.........J.....--
~ ~
o
Lower roller type stabilizer Ron.r type cutter
____-.tl;;,.~r------B~ body
Fig. 22.1.10. Schematic blind shaft drilling equipment setup.
drill pipe and down-hole drill tools are supported by the mast through a conventional crown and traveling blocklhook assembly. Static hook load capacities, for large-diameter blind-shaft drilling, may range from several hundred thousand pounds (kilograms) to more than a million pounds (half million kilogram) requiring more substantial masts than typically used for conventional rotary drilling. The rotary drilling motion is transferred from the rotary table to the drill pipe using a square section kelly bar. Mud is pumped to the down-hole system via a swivel located above the kelly.
/..'-
.,
1880
MINING ENGINEERING HANDBOOK Pump on " compressor
Cuttings and circulating f1uicrto pit
Pump on compressor
Cuttings and circulating fluid to pit
Drilled shaft
Cutters
Direct Circulation
Reverse Circulation
Compressor
Ground level
-~.m
Air, mud ........ and cuttings to pit II-'~r Fluid level
Fig. 22.1.11. Mud circulation systems for blind shaft drilling.
Air compressor Air, mud and cuttings to pit
Drilled shaft
- Mud
Dual string drill pipe Static water level
Air jets Cutters
Air Lift Circulation
Bit with plenim chamber Cutters,--~.=..!
Tool mud jets
Dual String Circulation
DOWNHOLE DRILLING TOOLS. These include the drill pipe and bottom-hole drilling assembly. Drill pipe is selected based on maximum tensile and torsional loading conditions and consideration ofmud circulation requirements; data sheets are available from the pipe manufacturers. A common US drill pipe, used for large-diameter blind shaft drilling, has an outside diameter of 13 3/8 in. (340 mm), weighs 90 Ib/ft (134 kg/m), and requires in excess of 100,000 ft-lb (136 kN-m) of makeup torque. The bottom hole drilling assembly includes a drill bit, mandrel, stabilizers, and donut weights. Cutters are mounted in cutter mounts or saddles and bolted to the underside of the flatbottomed drillbit. The drill bit is in turn bolted to the mandrel that serves as a base for locating donut weights. Donut weights are added to provide the required normal force at each cutter (typically from 10,000 to 20,000 Ib (44.5 to 89 kN) per cutter), which is a function of the relative hardness of the formation to be drilled. Donut weights are secured to the drill pipe by a holddown clamp which forms the top of the bottom-hole assernbly. Stabilizers may be added directly above the mandrel and toward the top of the bottom-hole assembly to assist directional control. OPERATIONS. An important element of blind shaft drilling is to maintain a straight, vertical alignment. The key to effective directional control involves minimizing the fraction of total effective bottom-hole assembly weight transferred to the bit while maintaining acceptable penetration rates. This maximizes the pendulum effect experienced by the bottom-hole assembly and, in conjunction with stabilizers, provides a straighter shaft. Most large-diameter drilled shafts use air-assisted reverse circulation. Drilling mud is added to the hole at ground level and circulated through the cutters and up the inside of the drill pipe. Air is added inside the drill pipe causing a density imbalance that induces flow rates sufficient to remove the drill cuttings. Air-assisted reverse circulation and alternative mud circulation systems have been described by Lackey (1982) for blind-shaft drilling at the Nevada Test Site (Fig. 22.1.11). The "dual-string airlift reverse circulation" method incorporates a
plenum chamber in the drillbit and dual string drill pipe. Mud and compressed air are pumped down the outer annulus of the drill pipe to the plenum chamber where the air separates from the mud. Mud flows through the plenum chamber and is forced through fluid jets located in the bit in order to clean the hole bottom. The air is routed through the top of the plenum chamber into the inside of the inner string and induces upward flow of mud and drill cuttings. Mud flows from the top of the drill pipe into the first mud pit where up to 90% of the cuttings drop out. Overflow from the first mud pit goes to the second and third where the remaining cuttings settle out. GROUND SUPPORT. A cake of mud is deposited on the shaft walls during drilling. The thickness and strength of this cake may be optimized based on cuttings removal and ground support requirements to prevent mud loss during drilling. Control of mud density and head (i.e., height of mud column) acting against this impermeable surface permits shaft excavation in poor ground. SHAFT LINING. The final lining for a blind drilled shaft typically consists of a ring-stiffened steel liner. This is equipped with external guides to facilitate grout line deployment and is outfitted internally. . Liner sections are fabricated offsite in lengths compatible with transportation and handling requirements; sections may be up to 60 ft (18 m) in length. The liner is lowered into the mudfilled hole using either casing jacks or the drill rig. Each liner section is aligned and welded to the one below to provide a completely water-tight membrane. Loads on the casing jacks may be limited, in the case of deeper shafts, by capping the bottom liner section so that the liner can be "floated" into place. Water is pumped into the casing to control buoyancy as the capped liner is lowered into the shaft. When liner installation is complete, the annulus between the steel liner and the shaft wall is filled with grout. Other lining systems, compatible with the concept of rapid, remotely controlled placement, have been developed but are not
RAPID EXCAVATION Table 22.1.3. Breakdown of Blind-drilling Project Costs Component
Percentage of Total Project Cost
Management Engineering and administration Drilling Casing Welding casing Cementing casing Site construction DriUing mud [)rillbit cutters and stabilizers Drillhole surveys Other (inc., radiographic inspection, crane services, tools)
2 8 25 24 10 7 3 7 6 1 7
in common use. These include slip forming (both bottom-up and top-down), jump forming, precast concrete cylinders, and remotely placed shotcrete. Finally, if ground conditions permit, conventional lining placement techniques (e.g., involving slip forming, jump forming, or shotcrete, placed using a galloway) may be used. BLIND-DRILLING SYSTEM PERFORMANCE PREDICTION. The performance ofa blind shaft drilling system is simply defined as a function of the operational penetration rate and the system utilization. Penetration rate is, in turn, a function of the geology (rock 'strength, fracture frequency, hardness, abrasivity); d~ll assembly (cutter type, size, and spacing; cutter load; and avatlable torque); and cuttings removal system. Moss et al. (1987) developed a drillability index to predict relative penetration rates in rock rated from exceptionally poor (Q = 0.001 [Barton et al., 1974]) to fair (Q = 10) with intact strengths ranging from 3000 to 43,000 psi (21 to 297 MPa). Variations in average penetration rates were smaller than expected and correlation with the index was poor. Several important observations were made as a result of this case study: 1. Lower than predicted rates of penetration in clay were thought to be due to plugging of the bit. Associated problems included reduced mud circulation rates and poor control of shaft verticality. 2. An increase in the rolling resistance when drilling in rock of lower rock mass quality was thought to result from fragments that were larger than those normally resulting from the cutting action. The results of this study serve to illustrate the potential shortcomings in generic performance prediction systems. The operational penetration rate for a blind-shaft drilling project can be estimated using available equipment specifications and the simple relationships suggested by Farmer and Glossop (1980) and Graham (1976). Adjustments are required for available thrust, calculated as 30% of the sum of the weights of downhole components corrected for buoyancy from the drilling mud and imperfect hole cleaning (Maurer, 1962). High utilization factors are possible for well-planned operations; for example, Hunter (1982) reports utilization factors of 74, 80, and 79% for the three Crown Point Project shafts. BLIND DRILLING COSTS. Table 22.1.3 provides a breakdown of blind-drilling cost components and their relative contribution to total project costs. It can readily be seen that the project costs are dominated by the acquisition and installation cost of the steel casing used for final lining.
22.1.4.2 Other Blind Boring Systems A manned blind-shaft boring (BSB) system, with operators located underground, was developed and demonstrated by the
1881
Robbins company in the late 1970s~ A 24.5-ft (7.5-m) diam~ter shaft was sunk to a depth of 587 ft (179 m), proving the appltcation of horizontal tunnel boring methods to vertical shaft boring. The BSB used a full-face rotary cutterhead equipped with 56, 13-in. (330-mm) disk cutters with a conveyor bucket elevator mucking system (Fig. 22.1.12). A second generation shaft boring machine (SBM) (Fig. 22.1.13) has subsequently been developed by a RedpathlRobbins team to mechanically excavate a 20- to 24-ft (6- to 7.3-m) diameter shaft. The second generation ~a chine incorporates a 10-ft (3-m) diameter cutter wheel fitted wIth 28, 15.5-in. (394-mm) cutters for a 20-ft (6-m) shaft diameter and a boom-mounted clam-type mucking unit. Hendricks (1985) presents a detailed prediction of the performance of the SBM in hard rock (18,000 to 30,000 psi or 124 to 207 MPa) shaft construction. Fig. 22.1.14 has been reproduced from this text to simply illustrate the predicted cost and schedule advantages of the SBM over conventional shaft mining methods.
22.1.4.3 Vertical or V-mole System The V-mole is a horizontal tunnel boring machine modified for vertical deployment by the German firm Wirth. First introduced to construct large diameter (16 to 21.5 ft, or 4.88 to 6.55 m) shafts in Europe in the early 1970s, it has since been used to construct four 23-ft (7-m) diameter shafts for an Alabama coal mine. Summary data for these case studies are presented in Table 22.1.4. The equipment, shown in Fig. 22.1.15, consists of the cutterhead, drive assembly, thrust and directional control cylinders, kelly, gripper assembly, and work platforms. . The gripper assembly, consisting of 8 to 12 grippers, prOVIdes resistance to the thrust and torque required for rock boring. Rotary motion is transmitted from the gripper assembly to the cutterhead through a kelly and up to 6 thrust cylinders are controlled by the operator to provide the required penetration rate. Muck is removed into a pilot hole by scrapers located on the cutterhead. The shaft lining is placed from work platforms located above the gripper assembly providing a continuous excavationllining cycle. Services and support equipment are deployed using techniques traditionally associated with conventional shaft sinking. In addition to the obvious differences in down-hole tools, the V-mole requires a pilot hole for muck removal. However, offset reaming can be controlled by the operator, allowing so~e deviation from pilot hole direction to be accommodated dunng sinking.
22.1.4.4 Raise Boring Systems Raise boring has been used to drill shafts ranging in inclination from horizontal to vertical with a majority of applications involving large-diameter holes steeper than 45° (see Table 22.1.5 for summarized case study data). System components and operational considerations are described below. A generalized raise boring equipment setup illustrating the available range ofdeployment methods is shown in Fig. 22.1.16. CONVENTIONAL RAISE BORER. Setup and Equipment-A raise collar is sometimes used to support the raise drill and provide sufficient vertical. clearance for the reaming head during holing through. ConventIonal shaft collar excavation and lining techniques typically are used to construct the raise collar. The raise drill is positioned on a steel substructure anchored to the collar lining after completion of the pilot hole. . The pilot hole can be drilled during mobilization of ~he major plant using rotary drilling methods, however, more ptlot holes are drilled with the raise drill after it is positioned. Since the
1882
MINING ENGINEERING HANDBOOK
m-il-t-lHJI _'-.a---=~v---+-Elevator discharge chutes
Cyclone separators --t-----i_____
Carousel paddles - - t - - - - L
Fig. 22.1.12. Blind-shalt-borer down-hole equipment (courtesy: Robbins Co., Seattle, WA, Model 2418B-189).
Moyno pumps ---tH=1i=--1If--tt-.-
Culterhead
~,.,-1l-if--- Main
bearing and seals
---i~1i~~I~U
Flight conveyers _ _ Bucket elevators
~~~~~~'~~~~~~~l'~
13' Disk cutter
-J
'------SCrappers
accuracy of this borehole directly influences shaft verticality, great care is taken to ensure that the pilot hole is drilled within owner-specified tolerances. Directional surveys are routinely performed every 25 or SO ft (7.5 to IS m) using conventional oil well survey techniques. A skid-mounted raise drill typically consists of a crosshead, positioned between two cylindrical guideposts, and hydraulic rams that lift the head and apply thrust, via the drill rods, to the reaming head. Rotary motion is provided through a ring gear and motor/reducer installed on the crosshead. The raise borer reaming head is transported underground during drilling of the pilot hole. Underground set-up involves assembly and attachment of the reaming head to the raise drill rods and preparation of the underground mucking system. Operations-Initial raise drilling is conducted at low thrust and RPM until all cutters on the reaming head are in contact with the rock. Thrust and rotation are varied by the raise operator to provide optimum penetration for each stratum encountered. Cuttings fall through the reaming head and are removed at the shaft bottom using the mine's mucking equipment; common practice involves maintaining the cutting pile flush with the mine roof to reduce airborne dust levels. The reaming head is immobilized and suspended from steel fixtures cast into the shaft collar concrete following break-
through. It can be removed from the shaft, after the raise drill and steel foundation are demobilized, using a small crane. A comprehensive paper by Worden (1985) provides a detailed, pragmatic description of activities involved in the reaming cycle. Ground Support-Raises are commonly unlined since raise boring is typically used in relatively competent formations. However, if ground conditions or use dictate the installation of a final lining, there are a few rapid lining systems to choose from. DOWN-REAMING RAISE BORING. A small proportion of raise-bored shafts have been excavated using an upward-drilled pilot hole with downward reaming to full shaft diameter. The advantage of reduced pulling capacity, associated with downreaming, is more than offset by problems associated with muck handling (muck must travel down the pilot hole alongside the drill pipe) and cutter replacement. DOUBLE-PASS RAISE BORING. Shaft size limitations in conventional raise boring are primarily associated with 'exponentially increasing torque requirements and the cost/feasibility of machine and tooling upgrades. As noted earlier in this chapter, the required torque is a function of the sum of the individual cutter rolling resistances multiplied by the mount radius and the torque required to overcome friction between the drill pipe and pilot borehole. As both the number of cutter kerfs and the aver-
- - - - - - - - - --_.-----
RAPID EXCAVATION
1883
••
Sinking bucket
--~---A-~./
Bottom deck of lining and equipping stage
Electrical swivel---...;....----.
Upper deck with hydraulic and electrical panels
Fig. 22.1.13. Shaft boring machine, 20 to 24 ft (6 to 7.3 m) (courtesy: Robbins Co., Seattle, WA).
Support columns
Slew rollers and drive ~:"""""~~~11f Slew assembly supporting: cutter wheel and drive
Mucking unit and clam bucket
""'--------- Support columns
age mount radius are proportional to the shaft diameter, the required cutterhead torque is proportional to the square of the shaft diameter. Excavation of shaft diameters beyond the singlepass capability (machine and drill-pipe capacity) of onsite equipment can be accomplished by reaming a smaller shaft with second-pass reaming to full size. Reaming heads should be selected to optimize the torque distribution and drilling load. Stabilizers are essential when second-pass reaming in longholes to prevent drill string whip. Alternatively, the raise may be sequentially reamed in short sections. TWO-STAGE SEQUENTIAL REAMING-HEAD RAISE BORER.
The two-stage sequential reaming head was first introduced in South Africa in 1985 as an alternative method of reaming largerdiameter, deeper shafts in hardrock. In operation, the smaller head is sumped in and advanced about 3 ft (1 m) (Fig. 22.1.17). This head is then retracted, and the remaining shaft area is reamed by the larger head; this cycle is repeated until the raise is completed. The first sequential-head raise borer, using an 8-ft (2.44-m) primary and 12-ft (3.66-m) secondary reamer, was used to bore three, 3oo-ft (91-m) deep ventilation raises at the Western Areas Goldmine. Wirth, in conjunction with Rocbor Raise-boring and Mining Contractors, subsequently developed the HG330
raise borer 14-ft (4.3-m) primary and 20-ft (6-m) secondary reaming head) that has been used during construction of raises up to 32oo-ft (975-m) deep (Schmidt and Fletcher, 1987). Many of the early problems, typically associated with a protypical method, have been resolved according to Schmidt and Fletcher (1987). Outstanding issues, traditionally associated with largediameter shaft construction, include excavation in poor quality and blocky ground, presence of large groundwater influx, and the impact of pilot hole accuracy on final shaft verticality. This latter constraint currently restricts most large-diameter raise developments to shafts that will not be outfitted. Blind Raise-boring-Blind raise boring, or boxhole drilling has been used in the South African goldfields to construct smalldiameter (5 to 6 ft [1.52 to 1.83 m]) raises up to 500 ft (152 m) in length. Raise boring can be conducted with a predrilled pilot hole or blind (without pilot hole) at advance rates between 4 and 6 ft (1.22 to 1.83 m)/hr (Friant et aI., 1985). Raise-boring System Performance Estimation-System performance, at a conceptual level, can be estimated using the case study data in Table 22.1.5 or by using Eq. 22.1.7 or 22.1.10. For example, a mine requires construction of 400-ft (122-m) long, 10ft (3-m) diameter ventilation shafts in granitic rock with uniaxial
1884
MINING ENGINEERING HANDBOOK compressive strength ranging from 24,000 to 30,000 psi (165 to 207 MPa). Excavation is in good quality rock, and stress or structurally controlled failures of the rock mass are not considered to be a problem for this shaft diameter. From the case study data, pilot-hole drilling rates can be expected to range from 6 to 9 fph (1.83 to 2.74 mlh), with reaming rates from 2 to 3.7 fph (0.61 to 1.13 mlh). The project appears to be well within the range of a number of raise borers, including the RBM-7, and will require a nominal 12.25- or 13.75-in. (311- or 349-mm) pilot hole; pipe size and tooling requirements should be selected by the contractor. Example 22.1.1. Estimate the penetration rate for a raise drill; using the following machine data available from the manufacturer,
Total Depth of Shaft (ft) 460°.-_ _ 50,-0_ _1_0,-00_ _1_50r-0_ _2_0r-00_ _2_5,...00_ _3_0-r-00------,
420 Ui' >ca ~
380
Q)
E
~
340
~
300
·uo c
o
()
ca 260 '5
I-
220
180 °
100
300
400 500 600 700 Total Depth of Shaft (m)
Total Depth of Shaft (ft) 1000 1500 2000
500
° Ui' c
200
800
900
2500
Maximum available thrust Maximum torque Maximum recommended thrust/cutter
1000
3000
Solution. Assume From Eq. 22.1.8,
6
~
g ~
P=
0" if/ 0"cl for
0.1 X 25,000 27,000
5
2,000,000 ft-Ib 2,150,000 ft-Ib 25,000 lbf
8,900 kN-m 8,474 kN-m 111 kN
granite = 1 : 14.
P=
3940 XlII 186 X 103
7ii 0
0.093 in./rev.
()
c:
·u0 2
7ii c:
PHR =
0
0.093 X 8 X 60 12
()
ca
2.35 mm/rev
4
PHR =
2.35 X 8 X 60 12
3
'5
3.7 fph, or
I-
2
0
100
200
300
400 500 600 700 Total Depth of Shaft (m)
800
900
1.13 m/h.
which must be derated for utilization Raise-bore Costs-Economic factors of mechanical raise boring have been discussed by Norman and Dye (1978), and a detailed breakdown of a raise-bore contractor's bid is presented in Nash (1982). Table 22.1.6 details the essential elements of a raise boring project for which costs should be estimated, and
1000
Fig. 22.1.14. Predicted performance of shaft boring machine (after Hendricks, 1985).
Table 22.1.4. Case Study for V-mole Shaft Construction Shaft Location (reference source)
Drilled Diameter
Depth
Advanced Rate
(in.)
(m)
(ft)
(m)
Excavation Duration (days)
Raine, 1934 (Table 1 in reference)
192 192 192 192 192 192 192 192 192
4.88 4.88 4.88 4.88 4.88 4.88 4.88 4.88 4.88
758 797 745 643 748 991 748 735 748
231 243 227 196 228 302 228 224 228
119 42 25 15 28 31 37 21 29
6.418.9 29.6 44.4 26.3 31.6 20.1 35.4 25.9
1.95 5.76 9.02 13.53 8.02 9.63 6.13 10.79 7.89
NA NA NA NA NA NA NA NA NA
NA NA NA NA NA NA NA NA NA
Sandy shale and sandstone (relative composition in Bruemmer, and Wollers, 1976, Table 1.)
(Table 2 in reference)
258 258 258
6.55 6.55 6.55
1532 1358 358
467 414 109
44 30 15
32.1 45.3 23.2
9.78 13.81 7.07
NA NA NA
NA NA NA
Sandy shale and sandstone
(Table 4 in reference)
276 276 276 276
7.01 7.01 7.01 7.01
1548 1978 1877 1929
472 603 572 588
67 69 32 32
23.1 28.7 59.0 62.2
7.04 8.75 17.98 18.96
105.8 95.3 122.9 107.0
32.25 29.05 37.45 32.61
Average
Maximum
(ft/day)
(m/day)
(ft/day)
(m/day)
Type Formation
Sandy shales, sandstones with 65 % quartz content, shales and coal seams. (Maximum UCS = 27,000 psi, or 186 MPa)
RAPID EXCAVATION
Ring beam platform
1885
Ring beams
Machinery deck
Gripper (8 pieces)
Fig. 22.1.15. Wirth V-mole-vertical mole (after Raine, 1984).
Kelly Cutterhead support (4 pieces)
cylinder (4 pieces)
~l.oo--+--Thrust
Gear
Hydraulic drive motors
Main bearings Cutterhead
Fig. 22.1.18 provides a rough guide to project costs as a function of raise diameter and required torque.
22.1.4.5 Selection of Shaft Construction Method Selection of the appropriate shaft construction technique for a given site involves an in-depth analysis of site geomechanical and geohydrological conditions, design criteria (e.g., diameter and depth, shape, use, life, etc), and availability and location of equipment, plus a determination of their relative impact on project cost and schedule. Cost is nominally the overriding consideration; however, timing may be critical, and higher shaft construction costs may be off-set by rapid access to the ore body. SITE-SPECIFIC DATA REQUIREMENTS. The geotechnical data set required for shaft design and construction bid package preparation is essentially the same for all shaft construction methods. The site investigation program should incorporate a fully logged core hole located within one shaft diameter of the proposed shaft centerline. (Note: a borehole on shaft centerline is preferred; however, the verticality of this borehole may impact the construction method if a pilot hole is required.) This borehole should be geotechnically logged (e.g., recovery, RQD, discontinuity description, lithology, rock description) and core samples selected for testing (e.g., uniaxial compression, slake durability, swelling, hardness, petrographic analysis). Samples may also be required by cutter manufacturers for proprietary drillability testing. Hydrogeologic data should be obtained by profiling the borehole using a down-hole (in line with drilling) or straddle packer test tool. Data analysis will provide the location and magnitude of groundwater inflows, shaft wall stability and rock
Disk cutters
support requirements, shaft liner design loads, rock mass data for blast design, and estimates of drill penetration rates, bit wear, cuttability, and overall suitability for mechanical excavation. These services are available from mining geotechnical companies and have been described extensively in the literature (e.g., see Hoek and Brown, 1980; Goodman, 1980). METHOD SELECTION GUIDELINES. As previously noted, the overall selection of an appropriate shaft construction method must be made on the basis of cost and schedule. This analysis may be supported initially by conceptual designs as described in this section. However, a final decision should be made in consultation with personnel experienced in the field application of each technique. Factors influencing the selection of each shaft construction method have been assembled and are presented in Table 22.1.7. At a broad conceptual level, the blind-shaft drilling method is preferred for conditions where 1. Freezing would be required for groundwater control during conventional shaft excavation. 2. Shaft lining requirements dictate the use of a fully hydrostatic. steel/concrete lining. 3. Rapid access outweighs added cost. 4. Disturbance of the surrounding rock is a prime criterion. 5. Access is not available for subsurface muck removal. Use of the blind shaft borer (BSB) or V-mole may be preferable where 1. Adverse impacts of groundwater inflow can be economically mitigated prior to construction (e.g., through grouting) and/or are relatively easy to handle during excavation.
1886
MINING ENGINEERING HANDBOOK Table 22.1.5. Case Study Data for Raise-drilled Shaft Construction
Raise Borer Reference, Model Number Harrison, 1972 Dynatec, 1989, RBM-7 Harrison, 1972 Harrison, 1972 Dynatec, 1989, RBM-7 '- Dyantec, 1989, RBM-7 Dynatec, 1989, SBM-1000 Woodward, E.M., 1983 Dynatec, 1989, RBM-7 Folwel, 1972 Dynatec, 1989, RBM-7 Dynatec, 1989, RBM-7 Dynatec, 1989, RBM-7 J.S. Redpath, 1989 J.S. Redpath, 1989 Dynatec, 1989, RBM-7 J.S. Redpath, 1989 J.S. Redpath, 1989 Woodward, E.M., 1983 Woodward, E.M., 1983 Dynatec, 1989, SBM-1000 J.S. Redpath, 1989 J.S. Redpath, 1989 Dynatec, 1989, SBM-1000 Dynatec, 1989, RFB-7 J.S. Redpath, 1989 J.S. Redpath, 1989 Dynatec, 1989, RBM-7 Dynatec, 1989, RBM-7 Dynatec, 1989, RBM-7 J.S. Redpath, 1989 J.S. Redpath, 1989 J.S. Redpath, 1989
ROB-81F (EMJ, 1981) J.S. Redpath, 1989 J.S. Redpath, 1989 Dynatec, 1989, SBM-1000 Dynatec, 1989, RBM-7 IR RBM-211 (Nash, 1982) IR RBM-211 (Nash, 1982)
Inclination (deg) 90 90 90 90 90 90 90 90 90 21.5 90 90 90 81 65 90 90 90 90 90 90 90 90 90 90 90 67 90 90 90 86.5 70 65 90 90 90 90 90 90 90 90 90 90 90
Rock Type Limestone
NA Quartz Diorite Dolomitic Limestone Granites Quartz, Biotite Arkose Granite NA NA NA Norite, Gabbro Quartz Diorite Fine-grained Calcite Fine-grained Calcite Granites Graintes Dolomitic Limestone Quartz Diorite, Porphyry NA Schist, Argillite Silicified Limestone Anorthositic Gabbro Anorthositic Gabbro Quartz, Biotite Schistose, Siliceous SiI Silicified Limestone Quartz Diorite Massive Limestone Salty Limestone Limestone, Shale Fine-grained Calcite Fine-grained Calcite NA NA Limestones, Sandstones Sedimentary Shales
UC Strength (ave) (psi)
Tensile Strength (ave) (psi)
Raise Depth (ft)
Machine Diameter (in.)
43,000 20,000 41,000 35,500 17,500 26,000 30,000 20,250 23,000 22,500 23,000 26,000 17,500 40,000 26,500 26,000 14,500 14,500 20,250 20,250 22,000 34,000 40,000 26,000 17,500 20,000 44,300 23,000 23,000 23,000 13,000 14,300 47,000 27,500 33,500 18,000 15,000 27,500 14,500 14,500 26,000 23,000 12,500 5,000
3,071 1,429 2,929 2,536 1,250 1,857 2,143 1,446 1,643 1,607 1,643 1,857 1,250 2,857 1,893 1,857 1,036 1,036 1,446 1,446 1,571 2,429 2,857 1,857 1,250 1,429 3,164 1,643 1,643 1,643 929 1,021 3,357 1,964 2,393 1,286 1,071 1,964 1,036 1,036 1,857 1,643 893 357
167 318 319 452 631 1,040 1,122 234 358 ·405 1,350 1,366 239 365 390 1,075 1,932 1,981 50 182 211 373 410 505 620 1,025 675 125 128 220 335 805 866 325 350 625 1,000 2,300 936 940 1,017 372 210 210
60 60 60 60 60 60 60 72 72 72 72 72 84 84 84 84 84 84 96 96 96 96 96 96 96 96 108 120 120 120 120 120 120 144 144 144 144 144 150 150 168 192 243 243
Drilling Time
Pilot Hole Diam (in.)
Pilot Hole (hr)
11.00
30
11.00 13.75 12.25 12.00 11.00 12.25 11.00 13.75 11.00 9.00 11.00 13.75 13.75 13.75 12.00 12.00 12.25 11.00 12.25 12.25 11.00 12.25 13.75 11.00 11.00 11.00 12.25 12.00 13.75 13.75 13.75 13.75 13.75 13.75 13.75 13.75 12.25 13.75 13.75 13.75
40 148 58 26 55 132 19 320.9 44.7 703.5 478.2 9.3 67.5 18 87.9 45.5 67 40 103.1 131.1 9 16 24 42 252.5 146
660 97.7 132.3 167 60
Ream (hr) 44 41 84 136 47 204 14 83 40 111 165 279 26 109.4 166.7 210 1,206 1,048 29 82.7 24 262.2 101.5 56 57 321.5 606.1 34 43 50 209.5 486.1 508 173 175 219 250 817 920 739 1,017 620 126 97
Pilot Rate (fph) 10.60
15.78 7.58 4.03 13.77 7.36 10.23 12.58 1.14 8.72 2.75 4.14 5.38 2.70 11.72 4.24 9.01 7.54 15.50 9.94 5.15 13.89 8.00 9.17 7.98 3.19 5.93
3.48 9.58 7.11 6.09 6.20
Reaming Rate (fph) 3.8 7.76 3.8 3.3 13.43 5.10 8.14 2.8 8.95 3.6 8.18 4.90 9.19 3.3 2.3 5.12 1.6 1.9 1.7 2.2 8.79 1.4 4.0 9.02 1;0.88 3.2 1.1 3.68 2.98 4.40 1.6 1.7 1.7 1.9 2.0 2.9 4.0 2.8 1.0 1.3 1.00 0.60 1.7 2.2
Conversion factors: 1 in. = 25.4 mm, 1 psi = 6.895 kPa, 1 ft = 0.3048 m.
2. Rock quality permits stand-up times compatible with the V-mole's mining cycle. 3. Rapid access outweighs added cost. 4. Minimum disturbance of the surrounding rock is a prime criterion. 5. Access is available for setup and underground muck removal. 6. Geologic structure permits pilot-hole drilling to the tolerances required by the shaft designer (V-mole only). 7. Immediate access to drilled strata for geologic logging, instrument installation, and testing is required. 8. There are no existing mine openings (BSB only). Raise drilling may be preferred where 1. Site conditions (e.g., rock quality and absence of large groundwater inflows) provide for stable excavation conditions. 2. Access is available for setup and underground muck removal. 3. Geologic structure permits pilot-hole drilling to the tolerances required by the shaft designer. 4. Design requirements (e.g., diameter, depth) are compatible with available equipment capabilities.
Under these conditions, raise drilling may offer a less costly alternative to all other methods of shaft construction.
22.1.5 RAPID EXCAVATION SYSTEMS FOR HORIZONTAL AND SUBHORIZONTAL MINE DEVELOPMENT Three rapid excavation systems are described in this segment, namely, full-face tunnel boring, mobile miners, and roadheaders or boom-type tunneling systems. Data are included to provide the reader with a means of comparing the relative merits of each system and to assist with method selection.
22.1.5.1 Full-face Tunnel Boring Systems Full-face boring systems or TBMs have been in common use in civil tunneling for many years but are used less' frequently in mining projects. Nevertheless, TBMs in European coal mines and the TBM at the Stillwater Mining .Company's platinum/ palladium mine near Nye, MT (Tilley, 1989) are proving the
RAPID EXCAVATION
1887
Fig. 22.1.16. Alternative raise boring methods (after Friant et al., 1985).
Pilot Hole Down
Upward Ream
Pilot Hole Up
Blind Drill Up
Ream Up
Downward Ream
Note: Vertical raises shown for clarity. Equipment available for most required raise inclinations.
viabilty of TBMs in mine development. These experiences may spur more widespread acceptance and utilization of this method of mechanical drivage within the industry. Constant development and improved tooling have resulted in machines that are capable of advancing large-diameter openings in strong igneous and metamorphic formations at rates that compete favorably, and in many cases exceed, conventional drill and blast methods. Tilley has reported advance rates up to 165 fpd (50 m/day) and 700 ft/wk (213 m/wk) for operations at the Stillwater mine. In addition to high advance rates, TBMs leave a smooth profile and minimize ground support requirements. A disadvantage of TBMs is their wide turning circle, although a range of mini-fullface machines are available that have smaller turning radii. The high initial cost of these machines is balanced by low running costs compared to drill and blast excavation systems. Fig. 22.1.19 shows the cutterhead, shields and operator console of a hard-rock double-shield machine designed for use in claystone, mudstone, sandstone and shale excavation. Fig. 22.1.20 shows a large-diameter hard-rock TBM equipped with a slotted shield to facilitate support installation. Breakthrough of ~ 20-£1 (6-m) diameter drivage in limestone is also shown. System components and operational considerations are described below. A generalized TBM equipment setup is shown in Fig. 22.1.21. SETUP AND EQUIPMENT. Full-face boring machines consist of a rotating cutting head. fitted with disk cutters, drag bits, button bits, or various combinations of these. Advanced machines are available on which the tool type can be changed and tool spacing varied. These developments have arisen from the need for machines that can cope with a variety of poor ground conditions. The cutting head may be an open structure with spoke-like cutting arms, or it may completely conceal the face except for muck-removal openings and access ways for tool maintenance. The open-type head gives better access to the face and tools, and can be used with a forepoling arrangement. Cutting forces are provided by the head rotation, while normal forces are provided by the thrust of the machine against the tunnel face. Reaction to this thrust is provided by grippers mounted on the TBM body, which in turn react against the tunnel sidewalls. Mucking is performed by buckets mounted on
the periphery of the cutter head, and muck is removed via a central conveyor system. OPERATIONS. Preparations for TBM excavation typically involve portal construction, placement of a concrete pad on which the TBM will be assembled, and installation of support services and equipment. Careful excavation, including the use of controlled blasting techniques, is usually required to mine the setup area to the tight tolerances required. Placement of a thin, 1- to 2-ft (0.3- to 0.6-m) concrete layer against the start-up face is recommended to reduce out-of-balance loads. TBM excavation is a continuous process, with cutting, mucking, and support installation proceeding concurrently. As the cutting head rotates, it moves forward, reacting against the grippers. The grippers are repositioned periodically when they reach the limit of their travel. On Wirth and larva TBMs, the grippers are also used to steer the machine. Robbins and Domag TBMs steer while boring using a floating main beam. Mucking in the immediate vicinity of the face is done by buckets located on the head periphery, and a central conveyor system that moves the muck through the body of the machine to a bridge conveyor. The bridge conveyor allows access for track laying and service installation without" disrupting the mucking operation. A variety of mucking systems can be used to haul muck to the surface, but typically conveyors or shuttle trains are used. Adequate muck removal rates are critical to optimum face advance rates. The orientation of the TBM is controlled by the grippers, in conjunction with a laser beam and microprocessor-controlled guidance system. These allow precise positioning of the machine, but problems may still be encountered in weak or soft ground in which the effectiveness of the grippers is greatly reduced. In these situations, 3-dimensional orientation control is facilitated by TBMs, that steer while boring. SUPPORT SERVICES AND EQUIPMENT. The trailing gear following the TBM provides an interface with the support utilities and equipment installed in the. completed tunnel up to 650 £1 (200 m) behind the machine. Components, and their configuration, are primarily a function of tunnel size and may include: 1. Bridge conveyor required to transport muck from the TBM to the muck cars.
1888
MINING ENGINEERING HANDBOOK
®
®
Fig. 22.1.17. Sequential reaming sequence (after Schmidt and Fletcher, 1987).
2. Dual-track rail system with remote muck car loading and handling (for larger diameter tunnels). 3. Telescopic ventilation line~ auxiliary fan(s), and scrubber system. . 4. Hydraulic power unit, transformer(s), and trailing cable. 5. Supply hoist for unloading and moving supplies. 6. Rock drill (for bolt installation), rock drill power unit, and rock bolt supplies storage. 7. Mechanical shop and cutter and supply storage area. System components are briefly described below; more detailed descriptions and specifications are contained in the references cited at the end of this chapter. Haulage System-Rail systems are commonly used to transport muck during TBM excavation. Various types of muck cars and trailing gear are available that facilitate continuous loading. Additional details, regarding conventional muck haulage systems, are presented in Chapter 9.3. Alternate muck removal systems, incorporating pneumatic or hydraulic transport of crushed muck, have been used in a small nUITlber of civil tunneling projects. Ventilation System-The ventilation system typically consists of 24- to 36-in. (610- to 914-mm) ventilation pipe with inline booster fans located several thousands of feet (meters) apart along the tunnel. The system is designed and deployed to maximize the cross-sectional area available for mucking equipment and is normally configured to exhaust air to the portal (Chapter 11.6). Electrical System-The electrical system typically consists of a high-voltage feeder cable with stepdown transformers mounted on the trailing gear and at strategic locations along the tunnel to service ventilation fans, lighting, pumps, etc. (Chapter 12.4). Total installed power requirements can be roughly approximated at twice the predicted TBM consumption. GROUND SUPPORT. Rock support requirements (Chapter 10.5) for hard-rock TBM drivages are generally minimal and estimates have suggested that the savings in rock support costs (compared to drill and blast) can offset the cost of the machine in as little as 4 miles (7 km) of tunnel drivage (McFeat-Smith, 1982). Hard-rock TBMs are commonly equipped with a partial or slotted shield, and when support is required, conventional rock support methods are used. Both soft-rock and hard-rock TBMs can be equipped with a full shield and segmental linings installed. This equipment enables hard-rock TBMs to cope with localized occurrences of soft ground. TBM PERFORMANCE PREDICTION-SIMPLIFIED ApPROACH. As noted in 22.1.3, simplified approaches to TBM performance prediction, such as those presented by Farmer and Glossop (1980) and Graham (1976), are preferred over methods that rely on detailed site data (e.g., as suggested by Lislerud, 1983) due to cost considerations. These simplified formulas are consistent with the conceptual level planning approach incorporated herein. However, the owner is well advised to maximize geotechnical data collection and interaction with the TBM manufacturer so that uncertainties in performance prediction. are minimal. Table 22.1.8 compares data available from one TBM manufacturer with cutting rates predicted using Eqs. 22.1.7b and 22.1.8b. System Utilization-TBM utilization is defined as the ratio of the productive TBM operating time to the total time available for tunnel drivage. An estimate of the machine cutting rate, derated for downtime using the system utilization factor, can therefore be used to provide an estimate of the project schedule. TBM system downtime has been analyzed extensively by Nelson et al. (1985); Tables 22.1.9 and 22.1.10 have been reproduced from this source to illustrate potential downtime sources,
1889
RAPID EXCAVATION Table 22.1.6. Raise-bore Project Cost Elements
Item
~
Estimated Quantity
Unit
Mobilization Site preparation Underground setup Production • Pilot bit cost • Reamer cutter cost • Operating labor • Supplies/maint.
Unit Price
Approximate Percentage of Total
Total Price
2-10 5-20
L.S. L.S. L.S.
2-8
ft ft ft ft
20-70 5-10 5-10
2-8
Note: High cutter cost is a function of rock properties (e.g., strength, hardness, abrasivity, bedding, etc.). Risks can be offset by requiring manufacturers to bid cutter cost on a per foot (meter) basis.
provide data for remediation, and selection of an appropriately equipped TBM cannot be overstressed. TBM PERFORMANCE PREDICTION. Example 22.1.2. The following data apply to the construction of a 24-ft (7.32-nl) diameter tunnel (Kerckhoff No. 2) in granitic rocks (o-e/ = 20,250 psi or 140 MPa; 0- if = 1450 psi or 10 MPa), 25% quartz content.
1500
1250
TBM Data (Robbins 243-217) Diameter No. Cutters/Diameter Thrust Torque Cutterhead Horsepower Cutterhead Speed
en
::::>
~
1000
~
Q) Q.
en0
0
750 -
24 ft 1 in. 57 X 15.5 in. 2,280,000 lbf 1,980,000 ft-Ibf 2200 hp
7.34 m 57 X 394 mm 10.14 MN 2.68 MN-m 1640 kW
5.8 rpm
5.8 rpm
Determine the overall rate in advance. Solution. Assuming operating thrust/cutter = 70% of machine capability,
Torque 160,000 ft Ib 500
Fe =
250
01.----_....L...--_.....J.....-_---L-_---L..-_----l...-_----l.._-...l._ _ 4 o 2 6 8 10 12 14 16
L..-------l
18
2,280,000 57
= 28,000Ibf/cutter
X 0.7
(124.5 kN/cutter)
Fe X 0.0158
1. Farmer and Glossup P
- - - - - = 0.306 in./rev
1450
Raise diameter, feet
(7.77 mm/rev)
Fig. 22.1.18. Raise costs as a function of diameter and torque (modified from Engineering and Mining Journal, Anon., 1981). Conversion factors: 1 ft = 0.3048 m, 1000 ft-Ib = 1.356 N-m.
0.306 X 5.8 X 60 12
= 8.9 fph
(2.71 m/h) average TBM utilization factors, and the relative impact of each major downtime component on utilization. One other major cause of downtime, excluded from the utilization calculation in Table 22.1.10, involves remedial work required when mining through major faults and shear zones. Major collapses and large inrushes of mud and water, heavily squeezing ground, and gouge that clogs the mucking system are a few of the reasons for significant delays when mining through tectonically altered zones. Case studies involving TBM excavation under adverse ground conditions have been reported by McFeat-Smith (1987) and are reported in summary form in Table 22.1.11. The importance ofa carefully executed geotechnical site investigation designed to evaluate potential impacts of these features and to
2. Graham method
Fe X 0.1
P
- - - - = 0.138 in./rev
20,250
(3.5 mm/rev) PHR
=
0.138 X 5.8 X 60 12
= 4.0 fph
(1.22 mlh) Assume utilization
=
40%.
PSYS = 1.6 to 3.6 fph (0.5 to 1.1 m/h), or 38 to 86 ft/24hr-day (12 to 26.4 m/day)
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MINING ENGINEERING HANDBOOK
Table 22.1.7. Comparison of Factors Influencing the Selection of Blind Drilling, Raise Drilling, and Conventional Shaft Construction Methods Factor Influencing Selection
Btind Shaft Borer (BSB)
Blind Drilling
Raise Drilling
Conventional Shaft Sinking
Vertical V-mole
Design considerations Safety
Utilizes underground operators working under cover and behind shields. No operation of large moving equipment required.
Does not require miners to work underground.
Only requires underground labor set-up and mucking. No labor required in the shaft prior to shaft lining.
System utilizes under. ground operators working in a controlled environment. Considered safer than conventional shaft sinking as operators are located remote from the working face.
Requires equipment operation in confined environment. Considered to be the most dangerous of the five shaft construction methods.
Shaft size
Depth is limited by skiphoist rope capacity.
Limited by required depth, available equipment, and cost. Drilled shaft diameters range from well size to 20 + ft.
Nominally limited by available machine torque. Short (300 ft) shafts have been raised at 20 ft diameter.
Limited by available equipment from 16 to 23 ft in diameter.
Required to be larger than 10-12 ft for most applications. Upper limit not controlled by method.
Shaft depth
Depth is limited by skiphoist rope capacity.
Limited by required diameter, available equipment and cost. See Table 22.1 .2 for case study data.
Raise depths up to 3200 ft have been reported.
Shaft verticality
Verticality can be controlled within extremely tight tolerances. Equal to or better than conventional.
Deviation can be estimated from geological data. Difficult to maintain an absolutely vertical shaft in sub-vertical structure.. Suggested tolerance for design purposes = 0.25-0.5°.
Shaft verticality is controlled by pilot borehole. Pilot hole accuracy controlled by directional survey and careful drilling practice. Design tolerance should be specified based on use requirements.
Shaft verticality is controlled by pilot hole. However, offset reaming can be controlled by the operator providing addi-. tional control when following a deviated pilot borehole.
Verticality can be controlled within extremely tight tolerances. The most accurate shaft construction method with regard to verticality.
Ground disturbance
Minimal mechanical disturbance of shaft wallrock.
Provides minimal mechanical disturbance of shaft wallrock.
Provides minimal mechanical disturbance of shaft wallrock.
Provides minimal mechanical disturbance of shaft wallrock.
Conventional (drill-andblast) excavation can resuit in deep seated blast damage.
Timing/schedule
Setup time is 1 month or more. Advance rate should be between 20 and 40 ft/ day. In general muck haulage and lining limit advance.
Typically much faster than conventional shaft construction methods. See Table 22.1.2 for case study excavation rates.
Reaming range from 15 fph (5 ft dia.), 3 fph (10 ft dia.), to 1 fph (15 ft dia.)
Reported rates between 30 fpd (16 ft dia.) and 60 fpd (23 ft dia.)
Generally restricted to one round per shift (e.g., 1 fph)
Groundwater
Same techniques as conventional sinking.
Method provides superior control of groundwater during excavation.
Groundwater controlled by pretreatment where necessary.
Large projected inflows require pretreatment (e.g., using grouting or freezing)
Large projected inflows require pretreatment (e.g., using grouting or freezing)
Support during excavation
Temporary or final support can be installed a short distance behind advancing face.
Support provided by hydraulic pressure and impermeable polymer skin permitting excavation in very poor ground conditions.
Not possible.
Temporary/final support can be installed a short distance behind advancing face.
Temporary/final support can be installed a short distance behind advancing face.
Final lining
Installed during excavation.
Steel/concrete composite lining typically used in weak ground. May incorporate bitumen layer for groundwater control and be designed for full hydrostatic, bituastatic or lithostatic loading conditions.
None usually required. Many rapid lining systems available-see text.
Final lining may be installed during excavation.
Final lining may be installed during excavation.
Requires existing underground access.
Requires existing underground access.
If a final steel lining is installed, considerable time can be saved through surface installation of guides, brackets and pipes. All furnishings can be aligned prior to welding to the downhole liner assembly.
Pilot hole deviation usually prevents raises from being used for man or materials winding.
Outfitting in-line with excavation and final lining if required.
Outfitting in-line with excavation and final lining if required.
-Initial
High
Medium
High
Low
-Operating
Medium
Medium
Medium
High
Operational considerations
Miscellaneous
Other considerations Shaft outfitting
Costs
RAPID EXCAVATION
Fig. 22.1.19. Model 1111-234 Robbins hard-rock double shield (courtesy: Robbins Co., Seattle, WA).
This compares to a forecast rate of 55 fpd (16.8 m/day) based on case study data analysis and an actual production rate of 60.5 fpd (18.4 m/day) (Woodward, 1983).
22.1.5.2 Mobile Miner A prototype of the Robbins mobile miner (Fig. 22.1.22) was introduced in 1984 for development ofa 3773-ft (1150-m) decline at M~. Isa mine, Australia. Advances of up to 12 ft/shift (3.66 m/shift) were made while mining a 12-ft (3.66-m) high, 21-ft (6.4-m) wide section in high-strength quartzite (16,000 to 39,000 psi or 110 to 269 MPa) (Boyd, 1987). Upgrades in dust control and sealing systems improved the initial utilization rates from 17 to 34%, with the best single eight-hour shift resulting in 12 ft (3.66 m) of drivage for 5.3 hours cutting time. Redesign of the cutter. wheel, to avoid high imbalanced loads and other upgrades resultm~.fro~ the ~t. Isa experience are anticipated to provide 50% utilizatIOn. This compares favorably with the 36% utilization required for break-even with drill and blast (Robbins, 1986).
22.1.5.3 Roadheader Systems Roadheaders have been in use in mining and tunneling for many years, and are known under a variety of names including
1891
Fig. 22.1.20. Large-diameter TBM and breakthrough of a 20-ft (6-m) diameter TBM drivage in limestone/shale (courtesy: Robbins Co., Seattle, WA).
boomheaders, boom-type tunneling machines and selective tunneling machines. Fig. 22.1.23 shows two of the many varieties of roadheader, and Fig. 22.1.24 illustrates commonly used terminology for the various machine components. Roadheaders were originally developed as a means of advancing roadways in underground coal mines, and early machines were limited to cutting relatively low-strength strata. Continuous development of these machines has greatly extended the range of applications, and they now are used in a wide variety of mining and civil tunneling work, including mine production (Sparks, 1980). Improvements in cutterhead design (Hurt et aI., 1982) and the increasing use of water-jet-assisted cutting (Barkam and Buchanan, 1987; Timko et aI., 1987; Hood, 1985) will result in a further extension of the range of roadheader applications in coming years. Roadheaders offer a number of advantages over full-face tunneling machines, chiefly related to flexibility. Roadheaders can cut a variety of cross sections, limited only by the basic dimensions of the machine, and are able to cut tight curves or corners. They are thus usable, for example, in room and pillar operations. Roadheaders can selectively cut narrow bands or
1892
MINING ENGINEERING ·HANDBOOK
Operator Console
-
Gripper Shoes
Disk Cutters
Fig. 22.1.21. Sectional view of tunnel boring machine.
beds and are thus suitable production mining tools if a careful sequence of mining and mucking is followed. Also they have lower initial costs than full-face machines. Roadheaders also otTer advantages over conventional drill and blast methods. One of the most important advantages of roadheaders is the avoidance of blast damage to the rock and the consequent savings in ground support costs. In addition, because mechanical excavation is a continuous process, shift time is more effectively utilized. Navin et aI. (1985) estimated that support requirements in openings excavated by a roadheader were reduced by 40% compared to drill and blast excavations, and advance rates by roadheader were 1.5 times as rapid as drill and blast. McFeat-Smith (1982) reported average utilization of about 50% in roadheader excavation systems compared to 33% for drill and blast. EQUIPMENT. The main components of roadheader excavation systems are discussed in some detail in this segment. However, because roadheader equipment is in a constant state of development, the reader is advised to consult manufacturers with regard to specific machines, detailed specifications, and applications. Roadheader cutting assemblies consist of a cutting head on a movable, hydraulically powered boom, mounted on a rotatable turret attached to a track-driven chassis. In addition to the cutting head, the machine also incorporates either a gathering arm or chain conveyor mucking system to remove broken rock from the face. The machine may be controlled by an-operator seated on the machine or located some distance away, perhaps inside a shield, or beneath a supported roof section. To improve machine stability during cutting, many machines are equipped with hydraulically powered stelling rams which are used to brace the machine otT the excavation sidewall. Pearse (1988) has studied the products of 13 roadheader manufacturers and tabulated basic specifications for about 60 machines, with total installed powers ranging from 70 to 600 kW, and weights ranging from 17 to 130 tons (15 to 120 t) (Table 22.1.12). Two types of cutting heads are available from several manufactures and are interchangeable on specific machines. These heads are termed transverse and in-line heads (Pearse, 1988), or
ripping and milling heads (Kogelmann and Schenck, 1982). Inline heads rotate co-axially with the boom, and arcing forces developed during transverse cutting may need to be resisted by stelling rams, particularly if the machine is of light weight. According to Pearse (1988), these heads are best suited to cutting rock with an unconfined compressive strength of 12,000 psi (80 MPa) or less. In-line heads require less thrust when sumping, and the head shape allows greater selectivity in cutting specific beds or bands, but stelling rams can be rendered ineffective if weak rock bands are present in the drift sidewalls. Transverse heads rotate at right angles to the boom axis, so arcing forces are resisted by the weight of the machine, and stelling is not usually required. An important advantage of this is. that machines can be lighter than similarly powered machines with an in-line head. Advances in cutter booms have resulted in machines equipped with telescopic booms, or booms with extended length for cutting high backs or crowns. Telescopic booms are useful for cutting on steep gradients or on weak floor materials where the thrust from the machine's travel system would not be adequate. In such cases, the machine can be stelled, and the thrust provided by the booms' telescoping hydraulics reacting against the machine. Kogelmann and Schenck (1982) have described soft-rock cutter booms (SRBs) and hard-rock cutter booms (HRBs), and report that the trend is toward the development of the latter. HRBs are reportedly more stable in cutting, producing less vibration and, consequently, less damage to bearings and picks. In low-strength formations, however, SRBs perform satisfactorily, and the increased cost of an HRB-equipped machine is not justified. Better cutter head design and analysis methods (Hurt and MacAndrew, 1981; Hurt et aI., 1982) can lead to improved pick utilization and pick life and reduced torque variation during cutting. These lead to reduced vibration and bearing damage. Cutterheads supplied with machines should be designed using sound engineering principles, and field modifications and repairs (e.g., replacing pick boxes) should be subject to the same controls and standards. Water-jet-assisted cutting heads are now available as standard or optional features on many roadheaders. Research in the application of low-, medium-, and high-pressure water jets
Phyllite, mica schist Phillite, mica schist Phyllite, mica schist Phyllite, mica schist Granite, granitic gneiss Diorite, greywacke, phyllite Dolomitic limestone Shale, sandstone, tillite Schist, gneiss Granite, gneiss Dolomite, layered shale Dolomite, layered shale Granite, schist Dolomite, phyllite Limestone, shale Shale, sandstone Shale, sandstone Dolomitic limestone, shale Dolomitic limestone, shale
Diorites and granites Granitic gneiss Granitic gneiss Granitic gneiss Granodiorite, aplite, tonalite Mica schist, quartz Dolomite, siliceous limestone Dolomite, siliceous limestone Sandstone, mudstone, coal
Phyllite, gneiss, amphibolite Quartzite, gneiss, dolomite Claystone, mudstone, shale Limestone, mudstone, sandstone Basalts Mica schist Mica schist
Rock Type
381,050 1,056,000 2,160,000 2,188,706 321 ,184 384,000 260,145 238,000 1,980,000 1,075,000 1,075,000 1,152,173 943,000 769,230 689,000 689,000 689,000 689,000 2,000,000
1,174,000 2,000,000 1,755,000 2,915,000 1,062,000 1,080,000 867,000 724,000 2,280,000 1,760,000 1,760,000 1,760,000 1,274,000 1,440,000 1,400,000 1,400,000 1,400,000 1,400,000 2,565,000
900 1750 1200 2420 800 800 600 400 2200 1400 1200 1200 900 1200 1050 1050 '1050 1050 2000 900 900 400 400 1200 1200 1200 1200 600 400 960 720 2400 2400
11.00 18.00 18.17 27.88 11.00 11.58 10.33 9.92 24.08 20.42 20.42 20.42 19.00 15.58 14.75 14.75 14.75 14.75 25.25 14.17 14.17 9.90 9.90 18.00 18.50 18.50 22.00 13.75 9.50 21.33 20.00 32.33 32.33
1264 1429 1521 1014 1429 1429 1571 1464 1446 1014 1357 1357 1000 1014 1518 1518 1518 1518 3018 1179 1571 1557 1429 1436 1757 1757 1375 1214 1536 893 893 1571 1571
17700 20000 21300 14200 20000
20000 22000 20500 20250
14200 19000
19000
14000
14200 21250 21250 21250 21250 42250
16500
22000 21800
20000 20100 24600
24600
19250 17000 21500 12500 12500 22000
22000
1,086,200 1,176,000 537,000 253,000 1,262,000 950,000 3,304,000 3,304,000
1,960,000 1,010,000 720,000 1,400,000 1,400,000 2,570,000 2,570,000
238,000 1,135,000 1,086,200
740,000 1,950,000 1,680,000 1,680,000
631,000 238,000
1,216,000 740,000
631,000
1,000,000
1,596,100
1200
17.38
1118
15650
1,216,000
411,9002517
1,397,700
900
11.67
836
64
46 34 24 46 43 64
38
24 42 42
31 24
34
36 35 35 35 35 57
41
44
44 44
27 26 24 57
27 40 39 61 26
36
15.5
15.5 14 14 14 14 15.5
15.5
14 15.5 15.5
15.5 14
15.5
15.5 15.5 15.5 15.5 15.5 15.5
14
15.5
15.5 15.5
15.5 14 14 15.5
17 17 17 17 17
17
17
11700
25
420,000
1,250,000
19
1000
24
11.58
400,000
1750
1,320,000
1000
10.50
Torque (Ib-ft)
24500
Thrust (Ib)
Cutterhead hp (hp)
Machine Diameter (ft)
Cutting Disk No. Dia. (in.)
643
Tensile Rock Strength (calc) (psi)
9000
Average Uniaxial Compressive Strength (psi)
0.32
50000
40156
42609 29706 30000 30435 32558 40156
44211
30833 46429 40000
39226 30833
35765
40000 40000 40000 40000 40000 45000
31073
40000
40000 40000
40000 33346 30167 40000
43481 50000 45000 47787 40846
44336
0.28
0.13
0.16 0.12 0.10 0.17 0.18 0.13
0.13
0.28 0.34 0.27 0.22 0.38 0.40 0.28
0.11 0.16 0.11
0.12 0.10
0.15
0.20 0.13 0.13 0.13 0.13 0.07
0.16
0.15
0.20 0.15
0.14 0.11 0.10 0.14
0.17 0.18 0.15 0.24 0.14
0.20
0.33
0.14
0.43
Graham (in.lrev)
0.24 0.36 0.25
0.28 0.22
0.34
0.44 0.29 0.29 0.29 0.29 0.16
0.34
0.33
0.44 0.33
0.31 0.23 0.23 0.31
0.38 0.39 0.33 0.52 0.32
0.44
0.74
0.95
55000
55908
Farmer & Glossop (in.lrev)
Max. Thrust/ Cutter (Ib/disk)
Estimated Penetration
Table 22.1.8. Summary Case Study for Tunnel Boring Machines
3.94
5.79 9.26 13.40 5.97 6.37 4.43
6.88
12.86 7.07 6.88
8.99 12.86
8.99
8.17 8.63 8.63 8.63 8.63 5.04
6.70
6.24
6.24 6.24
11.00 12.33 12.84 5.80
11.57 7.07 7.01 4.57 11.57
7.33
10.91
11.00
12.13
Cutterhead rpm (rpm)
2.52
4.49 5.67 6.55 5.09 5.81 2.83
4.33
6.94 5.72 3.92
5.61 6.37
6.82
8.06 5.69 5.69 5.69 5.69 1.88
5.21
4.60
6.15 4.60
7.70 6.54 6.61 4.01
9.96 6.19 5.18 5.38 8.28
7.27
18.26
7.86
25.95
Estd. Cutting Rate (Graham) (ft/hr)
5.57
9.92 12.53 14.48 11.25 12.84 6.26
9.58
15.35 12.65 8.66
12.40 14.08
15.08
17.82 12.58 12.58 12.58 12.58 4.16
11.52
10.16
13.60 10.16
17.02 14.46 14.62 8.87
22.01 13.69 11.46 11.90 18.30
16.07
17.37
57.37
Estd. Cutting Rate (f & g) (ft/hr)
5.7
4.8 5.87 7.47 6.94 7.55 5.9
11.1 4.72 9.04
7.87
11
9
9.84 7.61 7.78 7.64
10.8
7.7
6.53 8.2
7.57 5.22
11.5
23 8.9
12.2
12.1
Act. Average Rate (ft/hr)
12.5
10
5.3 18.9
9.1
9
30.2
Act. ' Best Rate (ft/hr)
(,)
CD
CC)
-L
oZ
<
» » -I
o><
c m
:!!
:JJ
»
= 25.4 mm, 1 ft = 0.3048 m, 1 psi
= 6.895 kPa, 1 Ib
=
28 40
700,000 1,850,000
310,000 1,050,000
500 900
10.67 19.00
2321 2075
32500 29050
34
766,700
2,702,900
800
14.96
1286
18000
69
3,510,000
2,760,000
2400
35.25
1500
21000
0.4536 kg, 1 hp
14 15.5
14
15.5
15.5
69
1321
3,~p,000
2,760,000
2400
35.33
1321
18500
15.5
69
3,510,000
14 14
27
Cutting Disk No. Dia. (in.)
27
18500
350,000
852,000 2,760,000
1786
25000 2400
350,000
852,000
35.33
1250 500
Torque (Ib-ft)
Thrust (Ib)
10.67
Cutterhead hp (hp) 500
Machine Diameter (ft) 10.67
Tensile ~ROCk Srength (calc) (psi)
17500
Conversion factors: 1 in.
Limestone, sandstone
Shale, limestone, siltstone Shale, limestone, siltstone Dolomitic limestone, shale Dolomitic Iimestone, shale Dolomitic limestone, shale Limestone, sandstone, shale
Rock Type
Average Uniaxial Compressive Strength (psi)
=
0.12 0.25
0.68
0.29
0.33
0.33
0.20
0.28
Farmer & Glossop (in./rev)
0.05 0.11
0.13
0.15
0.15
0.09
0.13
Graham (in.lrev)
Estimated Penetration
0.7457 kW.
25000 46250
79497
40000
40000
40000
31556
31556
Max. Thrust! Cutter (Ib/disk)
Table 22.1.8. Summary Case Study for Tunnel Boring Machines (cont.)
11.93 6.70
3.61
3.60
3.60
11.93
11.93
Cutterhead rpm (rpm)
3.21 3.74
2.41
2.73
2.73
5.27
7.53
Estd. Cutting Rate (Graham) (ft/hr)
7.11 8.26
5.33
6.03
6.03
11.66
16.66
Estd. Cutting Rate (f & g) (ft/hr)
7
6
5.1
5.2
6.75
6.75
Act. Average Rate (ft/hr)
10.4
7
10
Act. Best Rate (ft/hr)
.....
"
0 0
m
Z C
J>
%
Cl
Z
~
rn rn
eZ
Z
rn
Cl
Z Z
i:
~
U)
CD
RAPID EXCAVATION Table 22.1.9. Downtime Categories for TBM Operations Major Category
Individual Downtime Sources
TBM maintenance and repair
Cutterhead check and routine maintenance Lube oil system Hydraulic system Cutterhead motors Electrical system TMB conveyor Tunnel power supply Utility lines (air, water, fan line) Laser guidance and surveying Trailing floor conveyor Tripper or car pass Train delay-muck bound at heading Shaft or portal equipment Installing rail and switches Derailments Water inflow rock support installation- bolts and straps, steel sets Gripper/ sidewall support Equipment clearance Scaling loose material Muck jams in conveyor and hoppers
Backup system
Ground conditions
Cutter changes Other
Scheduled downtime Probe hole drilling Contractual downtime Moving TBM to new location Shift changes
Source: Nelson et al. (1985).
to assist cutting has indicated substantial improvements in pick life and dust suppression, and has reduced potential for frictional ignition of methane. In general, machines equipped with water jets can cut higher-strength formations than equivalent machines without water jets. Roadheader face mucking systems generally consist of an apron with either a scraper chain or gathering arm system (Fig. 22.1.25). Scraper chains generally extend around the perimeter of the machine, while gathering arms load muck onto a short conveyor that passes through the body of the machine~ In either case, the muck is generally discharged onto a bridge conveyor. From this point, the muck may be removed by a wide variety of methods including shuttle trains, conveyors, LHDs, trucks, and
1895
so on. This allows roadheaders to be used in conjunction with an existing mine haulage system, provided that adequate capacity exists. Without adequate capacity, the face will become muck bound and excavation will be delayed. In low-strength formations in which high cutting rates are possible, the mucking system may become the limiting factor in controlling advance rates. Roadheader cutting booms are generally mounted on a custom chassis with track propulsion. Track dimensions control ground pressures and should be carefully assessed in conditions in which the invert rocks are weak or prone to slurrying. Increasingly, booms are being mounted on a variety of other machines, such as hydraulic breakers, trucks, traveling gantries, and inside shields. Conventional manual/hydraulic control systems are being increasingly superceded by electronic/hydraulic control systems linked to microprocessor-based guidance and profile control systems. Guidance systems consist of fixed laser sources mounted some distance behind the roadheader and photoelectric targets mounted on the roadheader. Deviations of the roadheader from its desired position and orientation are detected by the targets, and corrections are automatically made. In addition to guidance control, automatic profile control is also available. These systems consist of a microprocessor programmed with the required excavation profile, and transducers that continually monitor the position of the cutting boom. The boom hydraulics are controlled electronically to ensure that the correct profile is cut. These developments enable very accurate alignment and profile control, which eliminates overbreak and ensures that the cutting sequence is optimum; work stoppages to allow time for survey work are also reduced. Electronic systems are also being used on roadheaders to monitor the condition of mechanical and hydraulic components, enabling preventative maintenance to be scheduled, reducing unplanned downtime. OPERATIONS. The excavation cycle commences with sumping or forming a cavity in the rock face to the operating depth of the cutterhead. During this operation, the cutter boom is kept stationary while the cutting head rotates and the whole machine is gradually moved forward on the tracks. When the sump has been formed, it is enlarged by tracking the cutting head across the face to the outer perimeter of the excavation. At the end of this arcing cut, the head is moved up or down, and another arcing cut is made across the face in the opposite direction to the first cut. Proceeding in this way, the opening is gradually enlarged to the full dimensions of the face. When the full cross
Table 22.1.10. TBM Excavation System Utilization and Downtime Percentagesa Downtime
Project
Tunnel Section
Utilization b
TBM Maintenance and Repair
Backup System
Ground Conditions"
Cutter Changes
Other C
1C0011
Outbound
39.2
18.3
19.2
16.7
2.8
3.3
1C0031
Inbound Outbound Inbound
45.1 35.4 35.1
8.8 18.8 26.0
27.4 18.7 17.2
10.5 19.3 14.7
2.0 2.2 2.7
6.2 5.6 4.3
Densmore and Goodman Legs East Heading
41.1
12.0
14.5
13.2
17.1
2.1
44.4
23.3
21.3
1.3
6.3
3.4
Average for All Projects
40.0
17.9
19.8
12.6
5.5
4.2
Culver Goodman TARP
Excluded Special Causes
Downtime Excluding Special Causes
Water inflow, stops at shafts and conrail Water inflow, stop at shaft Conrail Stop at shaft, relocate crane mucking Probe hole
60.8
Water inflow, mining past shafts, relocate TBM
55.6
54.9 64.6 64.9 58.9
60.0
Time percentages calculated excluding time before trailing floor assembly completed and excluding shift time required for special causes. TBM operating time, including time required for thrust cylinder reset. C Other includes downtime without explanation, shift changes, probe hole drilling, etc. Source: Nelson et al. (1985). a
b
1896
MINING ENGINEERING HANDBOOK Table 22.1.11. Case Histories of Tunnel Boring in Adverse Ground
Machine Type Tunnel Dia.
Geology
TBM (medium weight) 11.5 It
Moderately strong sandstones (Class 4)
TBM (heavy weight) 11.5 It TBM (heavy weight) 11.5 It
TBM (medium weight) 11.5 It
Geological Feature
Av. Rate Advance Delay to It/week Progress
75 It length of in· tensely jointed mud· stones and sand· stone Dolerite sill intrusion 1200 It of competent, into sedimentary se(Classes 2-3) quence 50,000 psi dolerite sill Pure mudstone at 60 It of soil infill zone roof level (Class 3) in roof (Class 5) overlying limecaused by dissolu· stone (Class 1) tion of limestone
465
Sandstone, mud· stones
390
Conversion factor: 1 ft
=
660 It throw fault giv. ing 50 It clay gouge zone with boulders (Class 5) and 33 It shattered zone (Class 4)
295
295
76%
Tunneling Problems in Adverse Zones
Design
Generally minorReliable mucking sys· Arch supports used tems and easy ac· cess for support installation 60% Slow cutting. Average Triple button disks progress reduced used although sinto 115 It/week. Cutgle disks may have ter costs very high. been better 342 hours Delays mainly for sup- Long roof shield pre· port and mucking vented installation due to collapse of of heavy temporary roof. support, conveyors chocked 67 hours Generally minorSuitable access for inArch supports stallation of arches used-timber packclose behind face. ing required for gripSingle gripper pads per pads most appropriate
Comments TBM well designed • for conditions Operating considered to be a success for such hard rock. Inadequate design features enhanced delays TBM very well designed for this severe condition
0.3048 m.
Fig. 22.1.22. Robbins mobile miner (courtesy: Robbins Co., Seattle, WA).
section has been excavated and the muck removed, a new sump is formed, and the entire process is repeated. The pattern of sumping and arcing cuts, relative to the direction of cutterhead rotation and geologic structure, influences the efficiency of roadheader excavation. SUPPORT SERVICES. In the vicinity of working roadheaders, high concentrations of airborne dust, generated during both cutting and transport of the muck, and high ambient temperatures commonly occur. In addition to exceeding statutory respirable dust limits, excessive dust may completely obscure the face, resulting in inefficient excavation and increased overbreak. High temperatures and humidity result in labor inefficiency and overheating of electrical motors. The use of waterjet-assisted cutting leads to a reduction in dust levels, but does not entirely eliminate the problem, and may actually increase humidity at the face. Meyeroltmanns (1982) has described practical methods of using ventilation to control airborne dust in the vicinity of roadheader faces, methods that also assist in controlling heat and humidity. GROUND SUPPORT. One ofthe primary advantages of selecting a roadheader excavation system over drill-and-blast methods is the elimination of blast damage and the consequent savings in rock support costs. McFeat-Smith (1982) estimates that in excavations requiring temporary support, the cost saving may be on the order of 10 to 15%, and in suitable ground, the need
(b)
Fig. 22.1.23. Roadheader-type tunneling machines. (a) Model RH 25 (courtesy: Anderson, Strathclyde) (b) Model ABM 330-1 (courtesy: Alpine Equipment Corp.). for support may be eliminated entirely. All types of rock support can be adopted for use in conjunction with roadheaders. However, because it can be relatively difficult to reverse a roadheader away from the face, the machine must be covered prior to shotcrete application. When ground conditions require, roadheaders
RAPID EXCAVATION Side View
1897
., f.. .;}j /1/ ~. ' .
Turret
Cutter Boom
x~ .~
Fig. 22.1.24. Typical roadheader showing main components. Conversion factor: 1 ft = 0.3048 m.
Apron
Track
can be mounted inside shields (Fig. 22.1.26) or advance within self-advancing powered supports (Fig. 22.1.27). The roadheader is supported on a slide mechanism and can move independently of the shield. Excavation, mucking, and erection of a segmental lining can proceed concurrently. The use of short shields allows tight turning circles to be maintained, although the system is less maneuverable than a non-shielded roadheader. Several cutter booms can be mounted in shields when large-diameter openings are required. ROADHEADER SYSTEM PERFORMANCE PREDICTION. Methods of predicting instantaneous and operational cutting rates for roadheaders were presented in 22.1.3. Overall system performance, assessed in terms of advance rate, is a function of OCR, face area, and utilization: Advance rate = OCR/Face Area X Utilization
(22.1.10)
Utilization is defined as the time available for advancing the face when all planned and unplanned machine stoppages have been accounted for. Machine stoppages generally fall into one of the following categories: 1. Planned maintenance of roadheader and backup equipment. 2. Unplanned maintenance of roadheader and backup equipment. 3. Mucking delays. 4. Ground control. a. Rock support installation. b. Control of water inflows. 5. Survey work. 6. Meal time/shift change time. Guidance for typical roadheader utilization has been suggested by Kogelmann (1988) based on the type of ground support to be installed:
Support Type None Rock bolts Shotcrete Shotcrete and rock bolts Steel sets Steel sets with full lagging
% of Cutting Time per Available Face Time
60 40 40 30 30 20 -
80 50 50 35 35 25
Examples of the application of cited performance prediction methods are presented below to further illustrate the process. Example 22.1.3. This example is based on results reported by Sandbak (1985) for mine drift development in quartz monzonite and dacite porphyry. The roadheader is a Dosco SL-I-20 with an 82-kW cutter motor. Table 22.1.13 summarizes key geotechnical properties for sections of drift over which cutting rates were recorded. Two prediction methods were used, and the results of these can be compared with observed OCR data given in the table. Solution. 1. Prediction Based on Method of Bilgin et al. (1988) The first step in the application of this method (see 22.1.3) is calculation of the rock mass cuttability index, RMCI (see Fig. 22.1.7). This is given as: RQD2/3 RMCI = UCS (kN/cm 2) X - lOO
(22.1.11 )
which is then used in the prediction equation to determine OCR: OCR
=
28.06 X 0.997 RMC1
(22.1.12)
Because the prediction equation given by Bilgin et aI., is for a somewhat more powerful machine than the Dosco SL-120, a slight linear correction is applied to Eq. 22.1.12 as follows: OCR = 28.06 X 0.997 RCM1 X HP/95
(22.1.13)
where HP is the head power of the Dosco SL-120. These results are in remarkably good agreement with observed OCR data. 2. Prediction Based on Method of Fowell and McFeatSmith (1976, 1977) The first step in this prediction is the calculation of specific energy SE required for cutting, given by: SE = -4.38
+
+
0.14(CI)2
0.000018(SN)3
+
+
3.3(UCS)I/3
0.0057(CC)3
(22.1.14)
where Cl is cone indenter hardness, DCS is unconfined compressive strength, SH is shore hardness, and CC is cementation coefficient.
1898
MINING ENGINEERING HANDBOOK Table 22.1.12. Typical Roadheader Specifications
Type
Wt
Head drive
Total power
Head type
Cut ht. max.
Cut width (one position)
kW
kW
liT
m
m
Ground pressure
Travel speed
Loading system
bar
m/min
(footnote)
Alpine Equipment Corp., PO Box 132, State College, PA 16804, US. ABM-40 ABM-110 ABM-132 ABM-160 ABM-200 ABM-300 ABM-400
15 24 34 45 47 70 90
40 76/110
132 160 200 300 400
90 190 200 370 410 500 600
liT liT liT liT liT liT liT
3.8 4.0 3.9 4.1 4.1 6.5 8.0
4.6 5.1 5.2 5.6 5.6 9.1 11.0
0.6 1.0 1.0 1.0 1.1 1.3 1.3
0.16 0.16 0.16 0.16 0.16 0.16 0.16
D/GA/SW D/GA/SW D/GA/SW D/GA/SW D/GA/SW D/GA/SW D/GA/Sw
6.8/13.6 6.8/13.6 2.6/2.8
GA GA GA GA GA
0.5 0.15 0-5.6 0-5.6
Anderson Strathclyde plc, 47 Broad Street, Glasgow G40 70W, Scotland. RH25 RH25L RH22 RH1/4 RH90
25.4 26 35 66 90
82 82 112 112
157 164 187 224 300
4.25 3.8
6.0 4.5
5/5.3
5.4/6
6.0 5.0
6.4 6.0
1.2 1.2 1.7 1.45 new machine
5.3 6.85 7.9 2.0
1.4 1.4 1.7 0-8.4
0.10
Atlas Copco-Eickhoff GmbH, Hunscheidtstr 154, D-4369 Bochum 1, FRG. ET-110 liT 25-30 110-132 185 ET-200 liT 40-45 160-200 340 ET-300 liT 80-90 200-300 460 ET-400 liT 100-110 300-400 560 Special components for mounting (see text) ETS-110 45* liT 110 75* ETS-200 160 liT ETB-110 22 110 liT t ETS-200 35 160 liT t * Depending on type of excavator. t Powered from other source.
4.0 4.7 5.3 6.3
11.0* 8.3 11.0* 8.3 to cut TBM tunnel ledges to cut TBM tunnel ledges
D/GA D/GA D/GA/FC D/GA/FC none none none none
Dosco Overseas Eng. Ltd., Ollerton Rd., Tuxford, Notts. NG22 OPO, England. MK IIA MD 1000 MD 1100 SL 120 MK liB LH 1300 LH1300(H) MK III TB2000 TB3000 TM1800 SB 400 SB 600
27.7 28.5 31.5 33 44 44 45.7 83 76 123
n/a n/a n/a
48.5 50 82 82 82 142 142 142 119X2 250X2 48.5 142 142
123.5 135 157 165 194 254 285 254 424 686 104 198 198
liT liT liT liT liT I I I I I I I I
4.1 4.2 4.2 4.1 6.0 4.1 4.1 6.0 3.3 6.0 5.2 dia. 4.7 dia. 5.8 dia.
3.0-5.8 2.7-5.7 2.7-5.7 2.0-4.3 3.0-7.4 3.2-5.6 3.5-6.0 4.0-7.1 4.0-7.7 4.5-8.9
1.5 1.2-1.6 1.4-1.7 1.5 1.2 1.5 1.5 1.4 1.9 2.2
4.7 7.2 7.2 13.8 8.4 10.1 9.2 5.4 9.6 12.8
n/a n/a n/a
n/a n/a n/a
EC SW/GA SW/GA GA SW/GA SW/GA SW/GA SW/GA SW/GA SW/GA various various various
Eimco (GB) Ltd., Team Valley, Gateshead, NE11 OSB, UK. TM Series: Approx. weight 100 t, 2 X 150 kW motors
GA
Herrenknecht GmbH, D-7635 Schwanan-Allmannsweier, FRG. SM2 SM1
n/a n/a
80 95
95 132
1.5-2.2.dia. 2.0-3.0 dia.
n/a n/a
n/a n/a
SC/BH SC/BH
1.0
2.2
nil/FC
Mannesmann-Demag, Buscherhofstr 10, 0-4000 Dusseldorf 13, FRG 68 H55 *Hydraulic excavator.
160
180 160
+
10
11-12.5
Mitsui Miil(e Co. Ltd., 1-1, 2-chome, Nihonbashi Muromachi, Chuo-ku, Tokyo, Japan. S50,S90, S100, S125
RAPID EXCAVATION
1899
Table 22.1.12. Typical Roadheader Specifications (cont.)
Type
Wt
Head drive
Total power
Head type
Cut ht. max.
Cut width (one position)
kW
kW
liT
m
m
Ground pressure
Travel speed
Loading system
bar
m/min
(footnote)
Paurat GmbH, PF 1220, D-4223 Voerde 2 (Friedrichsfeld), FRG. In U.K.: Dowty Mining Equipment Ltd., Tewkesbury, Glos. GL20 8HR. 44 43 70 110 120
E169 E195 E134 E200 E2428
140 170 230 350 300
I I I
225 263 353 512 480
2.3 4.2 3.05 6.0 7.5
lIT
I
3.4 5.2 4.1-6.6 7.6 8.9
1.45 1.5 1.7 1.8
5.2 6.2--6.9 7.5 7.5
1.5/1.8
5.4 4.5 17.6
FC GA FC FC GA
1.5 1.3-1.5 1.6 1.8
10 10 8 8
GA GA GA GA
9.3/18.6 9.3/18.6
Salzgitter Maschinenbau GmbH, PF 511640, D-3320 Salzgitter 51, FRG. STM STM STM STM
100 160 200 300
28 45 75 120
100 160 200 315
T T T T
200 282 330 509
4.0 4.2-5.0 5.3 6.1
TYAZHMASH, 26, 8. Serpukhovskaya UI., 113093 Moscow, USSR. 4PP-5
75
14-35m2 area
200
GA
Voest-Alpine AG, PF 2, A-4010 Linz, Austria. F-6A AM50 AM65 AM75 AM100
12 24 32-36 45-52 84-96
30-41 110 132-175 160-200 250-400
60-82 170 214-305 290-330 450-700
T T T T T
3.4-4.0 2.0-4.8 4.3-4.9 4.7-5.1 5.5--6.4
4.5 4.8 6.9 6.8-7.0 7.3-7.7
1.4 1.3 1.2-1.35 1.2-3.8 1.8-2.1
5/13/20
4-15 3-21
GA GA GA GA GA
70
T
3.8
4.1
1.0
10
FC
101 200
T T
4.3 4.1
5.2 6.0
0-95 1.0
26.7 10-31.5
FC FC
250 300 360 437 470
T T T T T
4.2 5.4 7.1 7.7 5.4
5.3 6.3 8.3 8.9 7.9
1.5 1.7 1.6 1.6 1.6
10-20 7.5-26.7 5.0 5.0 5.0
GA GA GA GA FC
5.0 6.0
Westfalia Lunen, D-4670 Lunen, FRG. 37 Fuchs WF-40 9 also WF-50 with 50 kW cutter motor 79 Dachs 53 13 Luchs 8-110 110 25 also N-110 and H-110 WAY 130 32 130 WAY 170 45 200 73 WAY 178 200 WAY 178/300 73 300 WAY 300 90 300 BH
=
Backhoe, D
=
Disc, FC
Source: Pearse, 1988. Conversion factors: 1 ft
=
=
Flight chain, GA
0.3048 m, 1 hp
=
=
Gathering arm, SC
0.7457 kW, 1 ton
=
+ 0.14 (0.0377 UCS + 0.254)2 + 3.30 UCSl;3 + 0.000018 (0.441 UCS - 8.73)3 + 0.0057CC3 (22.1.15)
SE = -4.38
Cementation coefficient is based on petrographic descriptions of the rock (McFeat-Snlith, 1977). When SE has been calculated, instantaneous cutting rate ICR can then be calculated from Eq. 22.1.5:
= HP/SE
Because the prediction equations use SE based on actual cutting time, a cutting time factor (CTF) correction must be
Scroll, SW
=
Star wheel.
0.9072 t.
Cone Indenter hardness and Shore hardness have been shown to be linear functions of unconfined compressive strength (Atkinson et aI., 1986), so that Eq. 22.1.14 can be rewritten in terms of USC and CC:
ICR
=
made to Eq. 22.1.5. CTFs for bulk excavations with an experienced operator are estimated to have values in the range 0.65 to 0.85, while for final trimming, it may drop to 0.3. Overall values for bulk excavation and final profiling of a face may be in the range of 0.45 to 0.65. In this example, 0.45 was assumed and applied as follows: OCR
= HP/SE X CTF or
(22.1.16)
= ICR X CTF
In this example the predicted values are low compared to the observed cutting rates, and this may be attributable to rock mass factors. Example 22.1.4. This example is based on results reported by Bilgin et al. (1988) for drivage of sewer tunnels in Turkey using a Herrenkneckt SM 1, and coal mine development drives using a Dosco MKIIA. Results are presented in Table 22.1.14.
1900
MINING ENGINEERING HANDBOOK
Gathering • Arm Loader For blocky, interlocked, wet and sticky materials. Effective loading on steep slopes.
Star • ,Wheel Loader For dry, non-interlocked and non-sticky materials. High loading rates at continuous flow. Low maintenance.
Spinner • Disk Loader Same as Star-Wheel Loader
Scraper· Conveyer Loader For non-blocky, non-abrasive materials Fig. 22.1.25. Roadheader loading (gathering) systems (after Kogelmann, 1988).
RAPID EXCAVATION
.",
1901
:.-
;.::;(':~; ..
•; :_~ ·f:..,1r,';':;~:
Fig. 22.1.26. Roadheader in horseshoe-shaped shield support sys-
tem (after Kogelmann, 1988).
Fig. 22.1.27. Model ABM-T road header with waling, hydraulic roof
support system (courtesy: Alpine Equipment Corp.). Results for both of the methods used are in good agreement with the observed results. Again, a CTF of 0.45 was assumed. Example 22.1.5. This example is based on cutting trials in British Coal Measures reported by Fowell and McFeat-Smith (1976), using a Dosco MKIIA. Although individual results (Table 22.1.15) show appreciable scatter, mean predicted values are in reasonably good agreement with the observed noncoal values. Again, a CTF of 0.45 was assumed. The correlation could be improved if suitable rock mass property corrections were ~ade. ROADHEADER PROJECT COSTS. Advance rates can be predicted based on the foregoing estimates of operational penetration rate and equipment utilization. Roadheader project costs are calculated based on these advance rates with loading of the following costs and resources:
Component
Costs Available in
Capitall1ease cost (1) ($/wk) Labor (operating and support) ($/hr) Cutters and maintenance (1) ($/yd 3, $/m 3 ) Mucking - equipment (1) ($/mo) - operators ($/hr) Ventilation - equipment (1) - installation and power Support - shotcrete ($/yd 3, $/m3 ) - concrete ($/yd3, $/m 3 ) - rock bolts (1) ($ each) ($ each) - steel sets (1) Notes: (1) Available from manufacturer.
and qualified labor, project schedule requirements and cost. The approach in this Handbook is considered to be applicable at a conceptual level of project planning. More detailed investigations and analysis should be undertaken using expert services available from geotechnical and design engineers, construction contractors, and equipment manufacturers prior to final method selection. SITE-SPECIFIC DATA REQUIREMENTS. Field data required for design and construction bidding are essentially the same for all tunnel construction methods. The site investigation program should include geologic mapping and fully logged coreho1e(s) along the tunnel alignment; in certain circumstances, horizontal boreholes are the most cost effective method of evaluating in situ conditions. Site investigation coreholes should be geologically and geotechnically logged (e.g., core recovery, RQD, discontinuity description, lithology, rock description, etc.) and core samples selected for testing (e.g., uniaxial and tensile strength, slake durability, swelling, hardness, etc.); samples will also be required by TBM manufacturers for proprietary drillability testing. Hydrogeologic data can be collected in-line with core drilling or after drilling has been completed. Data analysis will provide estimates of groundwater inflows and parameter values that input to the design of grouting programs; rock quality, strength testing, 'and in situ stress will be used to evaluate stability, design rock support or lining systems, and for blast design; laboratory strength and index test data will be used to estimate drill penetration rates, bit wear, cuttability, and suitability for mechanical excavation. METHOD SELECTION GUIDELINES. The choice of a tunneling method should be made using the approach outlined in 22.1.5 and the selection factors presented in Table 22.1.16. Mechanical mining systems are now available that can rapidly mine a broad range of rock types at gradients up to 25%, while negotiating relatively tight and variably curved alignments. Further innovations and developments in cutting technology, especially with regard to water and particle-assisted rock cutting; mucking systems (e.g., with regard to continuous mechanical conveying and hydraulic and pneumatic transport); and machine..fe.at.ures (e.g., access to cutterhead for maintenance, guidance systems, gripper and steering systems, etc.) are underway. However, motivation for these developments must come from a receptive and innovative mining industry.
Costs Calculated at ($/ft, $/m 3 ) ($/ft) ($/ft) ($/ft) ($/ft) ($/ft) ($/ft) ($/ft) ($/ft) ($/ft) ($/ft)
22.1.5.4 Selection of Tunnel Construction Method Selection of the appropriate tunnel/drift construction technique will involve an in-depth analysis of site geomechanical and geohydro10gic conditions, design criteria (e.g., diameter, length, shape, use, and life, etc.), availability and location of equipment
---_.
-- ----
----
MINING ENGINEERING HANDBOOK
1902
Table 22.1.13. Prediction of Operational Cutting Rates Using Data From Sandbak (1985) UCS Lithology
ROD
(MPa)
qm qm qm qm qm dp dp qm dp qm qm dp qm qm qm qm qm qm qm qm qm qm qm qm
72 70 70 35 58 38 18 22 39 39 55 18 56 52 52 63 63 52 50 50 7 47 47 44
200 170 170 120 116 172 100 95 90 92 85 90 115 130 130 157 157 130 145 145 70 115 115 120
Observed OCR (psi) 29,000 24,650 25,650 17,400 16,820 24,940 14,500 13,780 13,000 13,300 12,300 13,000 16,700 18,850 18,850 22,750 22,750 18,850 21,000 21,000 10,150 16,700 16,700 17,400
RMCI 353 294 294 131 177 198 70 76 106 108 125 63 172 185 185 253 253 185 201 201 26 153 153 153
Mean OCR
(m 3 /h)
(ft3 /hr)
Predicted OCRs (3) (2)
(1 )
5.9 11.9 5.1 13.6 17.8 17.8 21.7 11.9 13.2 17.0 20.8 14.9 13.2 17.4 11.9 13.2 9.3 12.3 12.3 12.7 12.7 14.4 12.7 14.0
208 420 180 480 629 629 766 420 466 600 735 526 466 614 420 466 328 434 434 449 449 508 449 494
142 84 196 120 80 75 90 162 133 103 80 134 110 80 117 86 122 113 108 105 176 106 120 109
13.7
484
109
53 34 78 44 35 22 33 62 58 45 39 52 47 32 46 33 47 45 40 39 72 43 49 43 43
Notes: (1) Using Bilgin et al., 1988. (2) Using Fowell and McFeat-Smith, 1976. (3) Predicted OCR, as 010 of observed qm = quartz monzonite dp = dacite porphyry CC = 8
REFERENCES Anon, 1981, "Raise Drilling," Engineering and Mining Journal, Vol. 182, No. 2, Feb., p. 96-102. Aleman, V.P., 1983, "Prediction of Cutting Rates for Boom Type Roadheaders," Tunnels and Tunneling, VoL 15, Jan., p. 23-25. Atkinson, T., Denby, B., and Cassapi, V.B., 1986, "Problems Associated with Rock Material Properties in Surface Mining Equipment Selection," Transactions Institution ofMining and Metallalurgy,. Sect. A: Mining Industry, Apr. Barkam, D.K., and Buchanan, D.J., 1987, "A Review of Water Jet Assisted Cutting Techniques for Rock and Coal Cutting Machines," Mining Engineer, Jul., pp. 6-14. Barton, N., Lien, R., and Lurde, J., 1974, "Engineering Classification of Rock Masses for the Design ofTunnel Support," Rock Mechanics, V01. 6, pp. 189-236. Bilgin, N., Segrek, T., and Shahriar, K., 1988, "Golden Horn Clean Up Contributes Valuable Data," Tunnels and Tunnelling, Jun., pp. 41-
44. Boyd, R.J., 1987, "Performance and Experimental Development of the Mobile Miner at Mount Isa," Proceedings Rapid Excavation and Tunneling Conference, Vol. 2, SME-AIME, New York, pp. 747768. Bruemmer, K., and Wollers, K., 1976, "Experience with Shaft Boring and New Developments in German Coal Mines," Proceedings Rapid Excavation and Tunneling Conference, SME-AIME, New York, pp. 126-1147. Crookston, R.B., Weiss, O.A., and Weakly, L.A., 1983, "Mechanical and Conventional Excavating Experience in Oil Shale Shafts and Tunnels," Proceedings Rapid Excavation and Tunneling Conference, Vol. 2, SME-AIME, New York, pp. 817-833. Farmer, I.W., and Glossop, N.H., 1980, "Mechanics of Disc Cutter Penetration," Tunnels and Tunnelling, Vol. 12, No. 6, Jul., pp. 2225.
Farmer, I.W., and Garritty, P., 1987, "Prediction of Roadheader Performance from Fracture Toughness Considerations," 6th International Congress on Rock Mechanics, Canada. Folwell, W.T., 1972, "Raise Borers Applied Horizontally," Proceedings Rapid Excavation and Tunneling Conference, SME-AIME, New York, Vol. 1, pp. 719-737. Fowell, R.J., and McFeat-Smith, I., 1976, "Cutting Performance of a Selective Tunnelling Machine," Tunnelling '76, Institution of Mining and Metallurgy, London. Goodman, R.E., 1980, Introduction to Rock Mechanics, Wiley, New York. Graham, P.C., 1976, "Rock Exploration for Machine Manufacturers," Proceedings Symposium on Exploration for Rock Engineering, Johannesburg, South Africa, Nov., pp. 173-180. Hanke, N.P., and Fabitzsch, R.B., 1983, "Technical and Economic Aspects of Rapid Shaft Construction with a V-Mole," Proceedings Rapid Excavation and Tunneling Conference, Vol. 2, SME-AIME, New York, pp. 1047-1066. Harrison, G.P., Green, N.E., and Bennett, W.E., 1972, "Some Aspects of the Art of Raise Boring," Proceedings Rapid Excavation and Tunnelling Conference, Vol. 2, SME-AIME, New York, pp. 11611183. Hendricks, R.S., 1985, "Development of a Mechanical Shaft Excavation System," Proceedings Rapid Excavation and Tunneling Conference, Vol. 2, SME-AIME, New York, pp. 1024-1045. Hoek, E., and Brown, G.T., 1980, Underground Excavations in Rock, Institution of Mining and Metallurgy, London. Hood, M., 1985, "Waterjet-Assisted Rock Cutting Systems-the Present State of the Art," International Journal ofMining Engineering, Vol. 3, pp. 91-111. Howarth, D.F., 1986, "Review of Rock Drillability and Borability Assessment Methods," Transactions Institution of Mining and Metallurgy, Vol. 95, London, Oct.
RAPID EXCAVATION
1903
Table 22.1.14. Prediction of Operational Cutting Rates Using Data From Bilgin et al. (1988). Example 22.1.4. UCS Location
ROD
(MPa)
1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25
70 22 50 50 50 50 80 50 60 60 60 60 25 25 65 30 30 30 30 21 75 75 75 75 75
104 150 83 70 73 76 156 157 82 94 126 128 159 146 106 103 89 154 133 55 30 23 43 32 16
Observed OCR (psi) 15,100 21,700 12,000 10,150 10,600 11,000 22,600 22,750 11,900 13,650 18,300 18,600 23,000 21,200 15,400 15,000 12,900 22,350 19,300 8,000 4,350 3,350 6,300 4,650 2,300
Predicted OCRs (3)
CC
RMCI
(m 3 /h)
(ft3 /hr)
(1 )
(2)
8 9 8 7 9 9 9 9 7 8 9 10 9 9 7 7 9 9 9 8 9 8 8 8 7
180 120 115 96 101 105 295 217 128 147 197 200 139 128 175 101 88 152 131 43 55 42 79 58 28
6.9 5.9 9.4 6.7 6.3 4.1 0.5 1.5 3.3 6.5 1.6 0.8 7.7 5.2 2.6 9.0 6.4 3.1 7.8 15.0 10.2 17.4 10.4 11.4 26.5
244 208 332 237 222 145 18 53 117 230 57 29 272 184 92 318 226 104 275 530 360 614 369 403 936
290 99 160 89 89 63 92 104 69 174 87 45 183 107 105 128 76 89 166 94 77 109 109 90 139
110 90 100 170 150 230 980 330 300 130 390 730 60 100 310 95 130 160 80 80 75 60 70 70 50
7.5
265
113
105
Mean OCR Notes: (1) Using Bilgin et al., 1988. (2) Using Fowell and McFeat-Smith, 1976 (3) Predicted OCR, as % of observed
Howarth, D.F., 1987, "Mechanical Rock Excavation-Assessment of Cuttability and Borability," Proceedings Rapid Excavation and Tunneling Conference, Vol. 1, SME-AIME, New York, pp. 145-164. Hunter, H.E., 1982, "Shaft Drilling-Crown Point Project," Proceedings 1st NMIMT Symposium on Mining Techniques, Shaft Sinking and Boring Techniques, Session Ill, Socorro, NM, May 7-8. Hurt, K.G., and Evans, I., 1981, "Point Attach Tools: on Evaluation of Function and Use for Rock Cutting," Mining Engineer, Mar., pp. 673-675. Hurt, K.G., and MacAndrew, K.M., 1981, "Designing Roadheader Cutting Heads," Mining Engineer, Sep., pp: 167-170. Hurt, K.G., Morriss, C.J., and MacAndrew, K.M., 1982, "The Design and Operations of Boom Tunnelling Machine Cutting Methods," 14th Canadian Rock Mechanics Symposium, May. Kogelmann, M.S., 1973, "Use of Boom-Type Miners," Proceedings 4th International Symposium on Salt, Northern Ohio Geological Society, May, pp. 471-482. Kogelmann, M.S., and Schenck, G.K., 1982, "Recent North American Advances in Boom-Type Tunnelling Machines," Proceedings Tunnelling '82, Institution of Mining and Metallurgy, London, May, pp. A155-A165. Kogelmann, W.J., 1988, "Roadheader Application and Selection Criteria," Alpine Equipment Corp., State College, PA. Lackey, D., 1982, "Blind Shaft Drilling," Proceedings 1st NMIMT Symposium on Mining Techniques, Shaft Sinking and Boring Techniques, Session Ill, Socorro, NM, May 7-8. Lislerud, A., et al., 1983, "Hard Rock Tunnel Boring," Norwegian Institute of Technology, Engineering Project Report 1-83, University of Trondheim, 159 pp. Lislerud, A., 1988, "Hard Rock Tunnel Boring: Prognosis and Costs," Tunneling and Underground Space Technology, Vol. 3, No. 1, pp. 9-17. McFeat-Smith, I., 1977, "Rock Property Testing for the Assessment of Tunnelling Machine Performance," Tunnels and Tunnelling, Mar.
McFeat-Smith, I., and Fowell, R.J., 1977, "Correlation of Rock Properties and the Cutting Performance of Tunnelling Machines," Conference on Rock Engineering, University of Newcastle-Upon-Tyne, Apr. McFeat-Smith, I., 1978, "Effective and Economic Excavation by Roadheaders," Tunnels and Tunnelling, Jan., pp. 43-44. McFeat-Smith, I., 1982, "Survey of Rock Tunnelling Machines Available for Mining Projects," Transactions Institution of Mining and Metallurgy, Sec. A: Mining Industry, Vol. 91, Jan., London, pp. A23-A31. McFeat-Smith, I., 1987, "Consideration for Mechanical Excavation of Rock Tunnels," Proceedings VI Australian Tunneling Conference, Melbourne, Australia, Mar. Meyeroltmanns, W., 1982, "Use of Ventilation Systems for Dust Suppression During Tunnel Construction with Roadheaders," Proceedings 3rd International Tunnelling Symposium-Tunnelling '82, Institution of Mining and Metallurgy, London. Moss, A., Zeni, D., and Hutchenson, D., 1987, "Prediction of Blind Hole Drilling Conditions: Geological Influences," CIM Bulletin, SO: 904, 33 p. Nash, W.R., 1982, "Raise Boring in Mining and Civil Applications," Proceedings 1st NMIMT Symposium on Mining Techniques, Shaft Sinking and Boring Techniques, Socorro, NM, May 7-8. Navin, S.J., Goff, J.S., and Moulton, W.W., 1985, "Roadheader Decline Development," Proceedings Rapid Excavation and Tunneling Conference, SME-AIME, New York, Vol. 2, pp. 751-770. Nelson, P.P., O'Rourke, T.D., and Glaser, S.D., 1985, "TBM System Downtime-Causes, Frequency and Durations on Six Tunnel Projects," Proceedings Rapid Excavation and Tunneling Conference, SME-AIME, New York, Vol. 2, pp. 847-866. Norman, N.E., and Dye, J., 1978, "Economic Factors of Mechanical Raise Boring," Transactions American Society of Mechanical Engineers, Vol. 100, Feb.
MINING ENGINEERING HANDBOOK
1904
Table 22.1.15. Prediction of Operational Cutting Rates Using Data 'from Fowell and McFeat-Smith (1976). Example 22.1.5. UCS Lithology
(MPa)
Coal Coal Coal Coal Coa.l Coal Coal Coal Coal Coal Coal L. Mudstone L. Mudstone L. Mudstone L. Mudstone M. Mudstone M. Mudstone M. Mudstone M. Mudstone M. Mudstone M. Mudstone M. Mudstone M. Mudstone M. Mudstone M. Mudstone M. Mudstone M. Mudstone U. Mudstone U. Mudstone U. Mudstone U. Mudstone U. Mudstone U. Mudstone U. Mudstone U. Mudstone U. Mudstone U. Mudstone U. Mudstone Mean OCR for Mean OCR for Mean OCR for Mean OCR for
24 24 24 24 24 24 24 24 24 24 24 34 34 34 34 25 25 25 25 25 25 25 25 25 25 25 25 40 40 40 40 40 40 40 40 40 40 40
Predicted OCRs (1)
Observed OCR (psi)
(m 3 /h)
3,500 3,500 3,500 3,500 3,500 3,500 3,500 3,500 3,500 3,500 3,500 4,900 4,900 4,900 4,900 3,600 3,600 3,600 3,600 3,600 3,600 3,600 3,600 3,600 3,600 3,600 3,600 5,800 5,800 5,800 5,800 5,800 5,800 5,800 5,800 5,800 5,800 5,800
10.3 9.4 8.6 14.0 9.8 57.2 17.2 25.7 32.4 64.8 25.7 10.0 7.2 7.8 11.6 6.8 1.2 6.8 5.9 6.2 7.8 6.9 14.3 18.4 26.7 18.6 29.9 4.9 1.6 7.4 4.6 3.3 6.9 10.6 8.6 23.5 2.1 1.2 25.0 9.2 12.5 6.8
Coal Lower Mudstone Middle Mudstone Upper Mudstone
3
(ft /hr)
(m 3 /h)
(ft 3 /hr)
364 332 304 494 347 2020 607 900 1140 2,300 910 353 254 275 410 240 42 240 208 220 275 244 505 650 943 657 1060 173 57 261 162 117 244 374 304 830 74 42 880 324 441 240
9.4 9.4 9.4 9.4 9.4 9.4 9.4 9.4 9.4 9.4 9.4 8.0 8.0 8.0 8.0 9.2 9.2 9.2 9.2 9.2 9.2 9.2 9.2 9.2 9.2 9.2 9.2 7.4 7.4 7.4 7.4 7.4 7.4 7.4 7.4 7.4 7.4 7.4 9.4 8.0 9.2 7.4
332 332 332 332 332 332 332 332 332 332 332 293 293 293 293 325 325 325 325 325 325 325 325 325 325 325 325 261 261 261 261 261 261 261 261 261 261 261 332 283 283 261
(2) 90 100 110 70 100 16 55 37 30 15 37 80 111 103 70 135 720 135 156 148 118 133 64 50 34 49 31 151 463 100 161 224 107 70 86 31 352 617 40 90 64 110
Notes: (1) Using Fowell and McFeat-Smith, 1976 (2) Predicted OCR, as % of observed CC = 8
Pearse, G., 1988, "Cutter Boom Tunnelling Machines," World Tunnelling, Mar., pp. 81-90. Raine, A.G., 1984, "Large Diameter Mine Shaft Construction," Stability in Underground Mining II, SME-AIME, New York, pp. 97-109. Robbins, R.J., 1986, "Future of Mechanical Excavation in Underground Mining," Mining Engineering. Roxborough, F.F., and Rispin, A., 1973, "The Mechanical Cutting Characteristics of the Lower Chalk," Tunnels and Tunnelling, Jan./ Feb. Roxborough, F.F., and Phillips, H.R., 1975, "Rock Excavation by Disc Cutter," International Journal of Rock Mechanics Mining Science, Vol. 12, pp. 361-366. Runge, D., and Zeni, J.T., 1987, "Blind Shaft Drilling's Application to the Establishment of a Gold Mine in Australia-A Case History," Proceedings Rapid Excavation and Tunneling Conference, Vol. 2, SME-AIME, New York, pp. 954-979. Sandbak, L.A., 1985, "Roadheader Drift Excavation and Geomechanica! Rock Classification at San Manuel, Arizona," Proceedings Rapid
Excavation and Tunneling Conference, Vol. 2, SME-AIME, New York pp. 902-916. Schmidt, N.F.B., and Fletcher, A.E.W., 1987, "Raiseboring Experience with the Wirth Two Stage Sequential Reaming Head and HG330 Raisebore," Proceedings VI Australian Tunneling Conference, Melbourne, Australia, Mar., pp. 307-317. Sparks, G.B., 1980, "Application of Hard Rock Continuous Miners to Cut-and-Fill Slot Stoping," American Mining Congress Journal, Vol. 66, May, pp. 29-33. Tilley, C.M., 1989, "Tunnel Boring at the Stillwater Mine, Nye, Montana," Proceedings Rapid Excavation and Tunneling Conference, SME-AIME, Littleton, CO, pp. 449-460. Timko, R.J., Johnson, B.V., and Thimons, E.D., 1987, "Water-Jet-Assisted Roadheaders," Proceedings Rapid Excavation Tunneling Conference, Vol. 2, SME-AIME, New York, pp. 769-781. Worden, E.P., 1985, "Raiseboring-The Reaming Cycle," Proceedings Rapid Excavation and Tunneling Conference, SME-AIME, New York, Vol. 2, pp. 929-955.
1905
RAPID EXCAVATION Table 22.1.16. Comparison of Factors Influencing the Selection of Horizontal Tunneling Method Factor Influencing Selection
Tunnel Boring Machine
DESIGN CONSIDERATIONS Miners work under sup• Safety ported grou nd. Special safety precautions may be required where access to the face of the machine is required for cutter changes. Method restricted to • Tunnel line circular shape. Tunand shape nels up to 35-ft diameter have been excavated in hard rock. • Tunnel length
Unsuited, due to capital equipment and mobilization costs to short drivages. TBM applications seldom economical for single drivages less than 7000 ft.
• Alignment
Clockwise rotation causes TBM to drift to left. Deviation is easily controlled, and alignment ensured through use of laser guidance system. Laser deviation may present problem in extremely long drivages. About 600 ft. Curves as tight as 300 ft are possible with modified equipment. About 250/0 limitation associated with mucking and gripping systems. Larger gradients are possible but require equipment modifications (e.g., hydraulic and lubrication systems). In good rock, TBM produces a smooth, hydraulically efficient bore. Overbreak and rockwall damage are practically eliminated.
• Minimum radius
• Normal maximum gradient
• Ground disturbance
• Timing/ schedule
Advance rates of 165 ft/ day and 700 ft/ week have been reported for minesized development drivages with moderate utilization « 35 % ).
Roadheader
Mobile Miner
Conventional Mining
Shields and hydraulic supports available for poor ground conditions. Equal safety potential to TBM, much safer than conventional mining.
Main application in hard, high strength rock requiring littleto-no support. Safety rated equivalent to TBM with improvement, to dust suppression system.
Inherently the least safe of all development mining alternatives.
Multiple heading levels permit a wide range of opening sizes and shapes to be mined. Single heading range is from 6-20 + ft. Non-circular tunnels up to about 2,000 ft (with sections larger than 200 sq. ft.) in softer sedimentary formation can be driven more economically using a roadheader. Accurate alignment and profile control available using laser guidance systems and microprocessor cutting boom control.
Flat-back, rectangular, minimum height controlled by size of cutter wheel (> 10ft).
Can produce full range of sizes/ shapes typically required.
No restriction.
No major physical restriction.
Minimum operating radius is about 40 ft. Alignment controlled by continuous survey or laser.
Easily controlled, no major restrictions.
25 ft.
65·ft (horizontal) 260 ft (vertical)
No major restrictions.
About 25 % controlled by equipment stability and mucking.
About 25 % controlled by equipment stability and mucking.
About 50 % controlled by equipment ingress/egress and mucking.
Minimal rockwall damage due to mining. Special precautions and careful operation required to avoid overcutting profile.
Minimal ground disturbance associated with mechanical cutting (cutterwheel fitted with disccutters).
Cutting rates in low strength rocks up to twice as fast as drilland-blast (see 22.1.3.5). Typically restricted to cutti ng rock < 15,000 psi.
Initial poor utilization (17 % ) improved to 31 % , 50 % utilization rates are achievable with advance rates (10' X 21' heading) on the order of 12 ft/shift.
Care required to minimize rockwall damage and overbreak. Recent advances in high-speed digital seismography allow accurate analysis of blast related rockwall damage and subsequent re-design to minimize effects. Typically the slowest of the four methods compared here. Advance rates typically equal to the smallest tunnel dimension per 8-hr shift for mine development headings.
1906
MINING ENGINEERING HANDBOOK Table 22.1.16. Comparison of Factors Influencing the Selection of Horizontal Tunneling Method (cont.)
Factor Influencing Selection
Tunnel Boring Machine
OPERATIONAL CONSIDERATIONS Large predicted in• Groundwater flows may require ground pretreatment (e.g., grouting). Groundwater control during excavation can result in significant reduction in utilization.
Roadheader
Mobile Miner
No major restrictions to groundwater handling as part of excavation cycle. Performance may be improved by pretreatment.
No restrictions to groundwater handling as part of excavation cycle.
Conventional Mining No restrictions to groundwater handling as part of excavation cycle. However, inflows typically slow progress and resuit in difficult mucking conditions. Wet conditions may require more expensive explosive types. Uneven profile and overbreak typically requires more rock support than used in mechanical excavation. Project costs may increase by up to 15 % unless careful blasting practices employed. Conventional drill-andblast can be used in practically all rock conditions. Preferred over TBM in mixed face conditions, where large variations in rock mass strength will be encountered, or when tunneling through rock masses containing faults and shear zones. No major restrictions.
• Ground support
Smooth, undamaged profile less likely to require ground support. Most conventional ground support systems can be installed in line with excavation.
Mechanical cutting provides smooth profile minimizing the need for ground support. Required ground support can be installed at the face.
Used in hard rock applications where ground support requirements are minimal. Ground support can be installed at face.
• Rock homogeneity
Cutting rate and cutter consumption are primarily a function of rocl{ type and properties. Best performance is in good quality, homogeneous rock. Poor quality and mixed face conditions can require careful planning and execution, and proper machine selection. High rock temperatures coupled with machine heat may require special cooling systems (e.g., air and water spray). Muck removal typically by rail, which usually limits advance rate. Continuous conveyors and pneumatic/hydraulic systerns may ensure a continuous mucking cycle, at required mucking rates, in the future. Auxiliary services housed on trailing gear (power, compressed air, water, ventilation, etc.). Power consumption high relative to other methods.
Care must be taken to ensure that variations in rock properties along proposed alignment are within the capabilities of the equipment.
No major restrictions. Current use in high strength quartzite.
No restrictions.
No major restrictions.
Muck discharge via conveyor at rear of machine. Rail and trackless haulage available. Moderately more versatile than TBM.
Same range of options as for roadheader.
Muck loaded from tunnel floor to truck or rail based haulage system, as part of excavation cycle. Scoop trams may be utilized for short haulage distances.
Similar to those required by TBM. AIthough installed power typically considerably less than for TBM, efficiency of converting power to rock breakage is less favorable.
Similar to those required for TBM. Improvements in dust suppression required based on initial field trials.
Power requirements considerably less than for mechanical systems. Drilling requires compressed air or water cooling.
• Rock temperature
• Muck removal
• Auxiliary services
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1907
RAPID EXCAVATION Table 22.1.16. Comparison of Factors Influencing the Selection of Horizontal Tunneling Method (cont.) Factor Influencing Selection
Tunnel Boring Machine
• Auxiliary equipment
Muck cars and train, roof support (if required).
Roadheader
OTHER CONSIDERATIONS • Costs High -Capital Medium -Operating Low -Support • Utilization 30-50% • General
Conversion factors: 1 ft
=
0.3048 m, 1 psi
=
Mobile Miner
Conventional Mining
Auxiliary equipment typically suited to other mine duties.
Auxiliary equipment typically suited to other mine duties.
Auxiliary equipment includes drill jumbo, loader, scoop-tram, truck or shuttle car or rail based haulage. Auxiliary equipment typically suited to other mine duties.
Low-Medium Medium Low
Medium Medium Low
Low High Potentially High
30-80% Capable of working in multi-face operations requiring cyclic excavation/support activities.
17-50% Improved utilization coupled with a relatively small turning radius suggest that boring machines of this type will gain favor for development in hard rock mining.
N/A Versatile equipment can be easily allocated to other mine construction/ ore mining operations.
6.895 kPa.
