Search Documents

By Helmut Hass, Michael S. Shultz

For many tunneling, deep shaft and deep excavation projects ground freezing is one of the best if not the only technique available to create a watertight and load bearing temporary support system in soft ground conditions. When appropriate attention is given to the details of the ground freezing system design, ground freezing is a very safe method and is economically viable with competing techniques and combinations of techniques. Recent innovative uses of the technique in Europe are highlighted in this paper and compared with current uses in North America.

Jan 1, 2005

By W. K. Hartley

INTRODUCTION AND HISTORY The Kimberley Division of De Beers Consolidated Mines Limited includes four underground mines in Kimberley viz. Dutoitspan, Bultfontein, De Beers and Wesselton Mine and two outside mines, viz. Finsch, an opencast operation and Koffie- fontein, an underground operation. Figure 1 is a map of the Kimberley district showing the location of these mines. Finsch is 150km west of Kimberley and Koffiefontein is 80km south of Kimberley. The head office of the De Beers Company is in Kimberley. The diamond sorting and trading companies for diamonds produced in South Africa are also in Kimberley. Diamonds were first discovered in the Kimberley area at Hopetown in 1866 on the banks of the Orange River. This, and later finds, triggered the start of the great "Diamond Rush". In 1870 the Jagersfontein (now closed) and Dutoitspan pipes were discovered, followed by those at Bultfontein, De Beers and Kimberley Mine (now closed) in 1871, Koffiefontein in 1880 and Wesselton in 1891. Finsch Mine was a more recent discovery in 1961. Finsch, due to go underground in 1986, and Koffiefontein, are fully self-contained with their own treatment plants, housing, services etc. The four Kimberley mines deliver ore by conveyor to a central treatment plant and use joint ser- vices. ,+lining is presently carried out mainly by Block Caving which has replaced the earlier Chambering method. De Beers Mine, on a limited scale, and Wesselton Mine have recently started to use Sub-Level Caving, and both Bultfontein and Du- toitspan are scheduled to use the same method at a later date. The combined ore tonnage produced by the Kimberley Mines is 17 400 tons per day, Koffiefontein 12 000 tons per day and Finsch Mine 17 000 tons per day. THE EVOLUTION OF MINING METHODS Initially, the open-cast mining was carried out to as great a depth as safety would permit. In the early stages of open-cast mining, the Kimberlite was removed in horizontal slices from levels 13m to 20m apart. The rock broken in the workings was raised to the surface by means of endless rope haulage inclines. As lower levels were attained these rope haulages were replaced by aerial hoists, inclined shaft haulages or inclined skip-ways. These methods were dependent upon the strength of the unsupported pipe walls, which became increasingly unstable with depth and a system was adopted whereby the ground mined in open excavations was tipped down passes to tunnels connecting with vertical hoisting shafts sunk in the country rock. As work in the open pit became more and more impracticable due to extensive falls of ground, this system was replaced by a combination of open-cast and underground mining, and eventually by wholly underground methods. Chambering With the advent of underground mining, a system known as "Chambering" (or "Inclined Chambering") was evolved. It was essentially a combination of shrinkage stoping and caving, whereby advantage was taken of the high pressure exerted by the loose rock that had accumulated in the open excavations due to the collapse of the pipe walls. This mining method was introduced around 1890 and was successfully employed, with little change, up to the early nineteen seventies. In this system each mine was served by a main hoisting shaft situated about 300m in the country rock. Main haulage cross cuts were driven from the shaft to the pipe at 200m (600ft) vertical intervals. A sub-vertical shaft, situated close to the pipe provided access to the working levels. Chamber levels were established from the sub-vertical shaft at 13m (40ft) intervals. See figure 2. On each of these levels successive 6,8m wide “cuts" were mined out across the pipe, from one rock contact to the other, and extended vertically to the footwall of the level above. Every 6,8m cut was made up of a chamber and a "back-pillart1, each approximately 3,401 wide. The chamber was mined by overhand shrinkage stoping. The levels were carried "en echelon", with each chamber on a level approximately three cuts ahead of the chamber on the level

Jan 1, 1981

By A. F. McKown, P. E. Sperry, W. L. Fourney, D. E. Thompson

INTRODUCTION Underground drill and blast excavation is usually accomplished with small diameter holes loaded with high energy explosives. The detonation pressures are extremely high and an extensive amount of energy is dissipated in the process; however, very little of this energy is used to create the specified fracture planes required for excavation. The available energy is expended in producing radially outgoing stress waves, in crushing the adjacent rock, and in producing a dense, randomly oriented radial crack pattern about the holes. This causes several undesirable effects: 1. Overbreak at the perimeter, which results in more muck to remove and extra concrete or shotcrete to fill the voids. 2. Damage to or loosening of the remaining rock at the perimeter which may result in the need for additional support measures. 3. Ground vibrations and/or airblast, which could result in damage to nearby structures and complaints by people living or working in the area. Several steps may be taken to minimize these effects. Where control of the perimeter is desired, smooth blasting techniques are generally employed in which the spacing and burden of perimeter holes are reduced and lighter, decoupled charges are used. Where vibration damage to structures is feared, control is generally exercised by limiting the charge weight per delay period so that the ground vibrations at the nearest structure do not exceed a peak particle velocity of two inches per second. Where blasting complaints may be a problem, con-

Jan 1, 1979

By R. R. Sayers

VENTILATION may be defined as the process by which vitiated air of an inclosed or partly inclosed space is continuously replaced by fresh air. Fresh air has been defined as invigorating pure air. Pure dry air at sea level contains the following gases:1 Analysis of Air at Sea Level PER CENT. Oxygen 20.94 Nitrogen 78.09 Carbon dioxide 0.03 Argon 0.94 Helium, krypton, neon, xenon, hydrogen, hydrogen peroxid, ammonia, ozone traces CARBON DIOXIDE LIMIT For many years carbon dioxide was used as an index of the purity of the air vitiated by the metabolism of people. It was, however, well known that the untoward effects produced on those exposed to such air was not due to the concentration of the carbon dioxide. It has been found that the untoward effects on the comfort and efficiency of those so exposed are due to the increased temperature and humidity. It has been found that men can breathe air containing many times the amount of carbon dioxide found in our worst ventilated theaters and assembly halls, which, according to Rosenau, do not contain above 0.5 per cent. carbon dioxide.

Jan 7, 1926

By D. T. Farnham

The shale used at the Renton plant of the Denny-Renton Clay and Coal Co. for the manufacture of vitrified paving brick occurs in a hill rising from 200 to 300 ft. above the level of the valley in which the plant is situated.

Jan 1, 1915

By Shibata Y, Shimizu R, Fukushima A

Kobe is divided by the Rokko Mountains, runn- ing from east to west. The southern section of Kobe is highly developed in contrast with the northern section, but big, new residential pro- jects like the Hokusetsu New Town and Hokushin New Town are steadily being developed. To handle future traffic between the two sections of the city, the double-tracked Hokushin Line, now under construction, will connect the Shin Kobe Station of the Kobe Municipal Transit (Kobe's subway system) to the Tanigami Station of the Kobe Electric Railway. The locations of Kobe city and Hokushin Line are shown in Fig.-1 and Fig.-2 respectively.

Jan 1, 1987

By D. W. Loucks

There is plenty of evidence to indi-cate that at least one of man's chief interests in life is to make himself as comfortable as possible. If you doubt this, just watch the fellow next to you for the next half hour trying to find the most comfortable position that a hard chair has to offer. Comfort, however, does not always mean an easy chair. To some, it may mean a wealth of money; to another, freedom from worry. But to most of us, it means first of all a comfortable atmosphere in which to live, and to a great many of us it probably also means freedom from that annoying task of firing the furnace. Today more than ever before. automatic heat is one improvement that is placed high on everyone's list. Perhaps this is because automatic heating is becoming relatively cheaper. Perhaps it is because of a good publicity campaign on the part of the oil and gas men or maybe it is just that we are getting lazier day by day. At any rate, almost every issue of Better Homes and Gardens, House Beautiful, or your other favorite home magazine carries an article extolling the virtues of this or that automatic heating system. If I were to ask you to name the first thing that came to your mind when I said automatic heat, you would prob-ably say either gas furnace or oil burner. Or if you had just been studying heating systems, you might possibly say heat pump. But chances are you would not mention anything about coal, and yet coal is the most common source of the greatest automatic heat of them all. I say this because coal is the fuel used almost universally by the district heating industry in producing and delivering to certain heavily populated areas heat ready to use at the touch of a valve or the click of a thermostat. Although the industry is over a half century old, it has not experienced the widespread development of other utility industries because of certain limitations which I believe you will realize from the next few minutes discussion. District Heating Operations We may define district heating as any operation where two or more buildings are heated from a central heating plant. The method of heat transfer may be hot water or in some cases warm air, but generally the medium of heat transfer is steam. So universally is steam used that the industry is frequently referred to as the district steam industry. The Allegheny County Steam Heating Co. which operates the district heating system in downtown Pittsburgh is a subsidiary of the Du-quesne Light Co. Although organized in 1912 primarily as a means of securing the electric load of downtown buildings, the service has now become so valuable and so popular that it is no longer considered a necessary adjunct to the electric business but rather a separate business standing on its own feet. Fig 1 shows the layout of the plants and distribution system of downtown Pittsburgh. Two generating plants, one known as the Stanwix and the other as Twelfth Street, supply the area. Each has two boilers with capacity totaling 1,350,000 lb per hour. The Stanwix Plant is supplied coal by truck. The coal is pulverized at the plant and burned as powdered fuel. Coal is supplied to the Twelfth Street Plant also by truck but the boilers arc stoker fired. Over 1 1/2 miles of tunnel house a portion of our main lines, but it requires over twelve miles of pipeline, ranging in size from 32 down to 1 in. in diameter, to supply all our customers. The distribution system consists of two systems in a sense, one high and one low pressure with certain interconnections between the two. Our high pressure system supplies steam up to 125 Ib to some but not all customers, while the low pressure system operates in the range of 10 to 20 psi. Note that the two plants are tied together through large steam mains and that the system to some extent is a loop system, making it possible to have a portion of the line shut, down without interrupting service to any customer. Fig 2 conveys a picture of the extent to which steam service is used in the downtown triangle. The black area indicates the buildings which now use district steam. The dotted area indi-

Jan 1, 1950

By Francis S. Kendorski

INTRODUCTION Drift design problems in caving operations are a re¬sult of the geologic factors contributing to the overall success of the system, of the engineering factors dictated by economic and technical considerations, and of ore production practices. Combining these factors, a rational underground sup¬port system of rock reinforcement, light steel channel section or welded wire fabric, and shotcrete can be de¬signed based on rock fracturing, rock load, abutment loadings, ground movement, expected repair and desired flexibility. The design concept uses the effect achieved by restraining, reinforcing, and maintaining some of the intrinsic strength of the fractured rock mass composed of interlocked blocks of intact rock and rock fractures. Three different examples of drift support design in hypo¬thetical mines using the caving system are given. Caving is a system of underground mining where ore is extracted by means of gravity after the ore body is allowed to fail by removing support from underneath. The rock mass of the ore body fractures and flows ver¬tically downward to let gravity do as much work as pos¬sible. Caving differs from many other mining systems in that blasting is used only to initiate the rock mass failure by removing the rock supporting the ore but not to break the ore itself. The initial movement of the rock mass dur¬ing failure and the consequent crushing and grinding during the continued movement serve to reduce the ore to particles of a manageable size, with only limited sec¬ondary blasting necessary. The broken ore is extracted from the bottom of the failed rock mass through funnels of some sort pre-excavated in the rock. Ore extraction must continue or the swell of the broken rock will even¬tually fill the cavity and stop further rock mass failure and movement. The excellent general discussion on block caving in the SME Mining Engineering Handbook (Julin and Tobie, 1973) adequately covers the principles and application of this type of underground mining. Many rock mechanics aspects of block caving have been covered by others (McMahon and Kendrick, 1969; Swaisgood, et al., 1972; Mahtab and Dixon, 1975; King, 1946) and will not be reviewed further. Maintaining the stability of production drifts is one of the most troublesome problems plaguing the mine manager in a caving operation. Many factors contrib¬ute to drift support problems, and identifying the causes of instability and producing a reasonable support design are two steps toward achieving stability consistent with the mine plan. This chapter sets forth a technique for the design of support systems for production drifts in caving opera¬tions. The basic support system elements employed are rock reinforcement, welded wire fabric, and shotcrete. Recognized as contributing to the design are the factors of rock load, additional load from mining activity, rock fracture characteristics, repair expected, and flexibility. It must be emphasized that the drift must first be stabilized as for a tunnel, and the additional strengthen¬ing for mining-induced loads cannot contribute to the initial premining stabilization, or the reserve of strength is used up. MINE PLANNING The efficient mine planning engineer not only must satisfy the economic, human, and environmental aspects of his task but must also consider the mechanical con¬sequences of his plan. The problems created for the mine by placing parallel drifts too close, by crossing drifts on different levels with inadequate, if any, separa¬tion, and by installing connections and crossovers in the haulage plan without regard for the effective spans cre¬ated, are only a few of the problems a mine planning engineer can create for himself and the mine. The effect on immediately adjacent mine areas when an area is caved is important to drift design because the removal of vertical support from a rock mass causes the weight of that rock mass to be shifted elsewhere. The adjacent rock mass will carry this load and reach a new equilibrium with the applied stress. The advancing front of stress increase that results from caving (and many other mining systems) is generally called the abutment load and is the increase in stress over the gravity or tec¬tonic stress that already exists, as shown in Fig. 1. In general, the abutment load will be similar in nature to the stress change found around an opening in rock and will be taken as causing an increase in the vertical prin¬cipal stress, due to an approaching caving boundary, so that

Jan 1, 1982

By D. E. Spiller

Introduction Precious metals recovery from ore can generally be accomplished using gravity concentration, flotation, and/or hydrometallurgical (leaching) techniques. The objective of this paper is to show why gravity concentration can be an important part of recovery systems that employ leaching as the primary unit operation. A brief discussion of modem gravity concentration equipment is also presented. Discussion Gravity concentration of ores has generated increasing interest in recent years. Reasons for this interest include: • Gravity concentration is environmentally attractive. There is little or no use of reagents. Hence, it is relatively nonpolluting. • The cost of cyanide has continued to increase. Therefore, cost savings may be realized whenever leaching feed tonnage can be reduced by preconcentration. • Compared to flotation and leaching, gravity equipment costs are low per processed ton. Field installation costs for gravity circuits usually are less because many' units are supplied as self-contained modules. Also, the cost required to supply services, particularly power, to a gravity plant site are also less. In situations where preconcentration at coarse particle size is applicable, significant grinding equipment savings may be possible. • Gravity circuit operating costs are also relatively low compared to typical flotation and leaching circuits. Reagents, power, maintenance, and manpower savings in a well-engineered gravity plant may be realized. Again, if grinding is reduced, significant power and steel (media and liners) savings are possible. •In recent years, more efficient gravity concentrating devices have been developed. Benefits to Precious Metal Leaching Gravity beneficiation can complement precious metal leaching in two ways. First, the recovery of coarse liberated values before leaching may reduce leach time requirements and may reduce reagent consumption. Second, gravity preconcentration can reduce the size of a leach plant by decreasing the quantity of material to be leached. Coarse gold and silver have been shown to leach rather slowly. Kameda (1949) and Habashi (1967) have investigated the kinetics of cyanide leaching systems. They concur that in a heterogeneous reaction, the rate of gold and silver dissolution is directly proportional to the surface area. Thus, the instantaneous rate of dissolution for spherical 0.37 mm (400 mesh) gold is theoretically -25 times faster than for the same amount of gold at .841 mm (20 mesh), based on data from Fuerstenau, Chander, and Abouzeid (1979). Conversely, coarse liberated, +.841 mm (20 mesh), gold is more readily recovered by gravity concentration than is fine, -.037 mm (400 mesh) gold. Therefore, it is apparent that the two recovery systems complement one another. Figure 1 data demonstrates the potential synergism. A sample of - 3.327 mm (6 mesh) Nevada gold-bearing ore was cyanide leached using conventional bottle-roll test procedures. Gold extraction was determined as a function of leaching time. A second sample split from the same leach feed material was hand jigged to remove a coarse heavy mineral fraction, including virtually all of the +.210 mm (65 mesh) liberated free gold. This second sample, with the coarse gold and heavy minerals removed, was subjected to an identical cyanide leach procedure. Figure 1 presents the resulting comparative extraction data. Note that the percent gold extraction for the sample containing no +.210 mm (65 mesh) free gold includes the coarse gold recovered by gravity. The data show that the sample containing coarse gold required about 72 hours of leaching time to achieve 80% extraction. This compared to about 22 hours of leaching time for 80% gold recovery from the sample that contained only -.210 mm (65 mesh) free gold. Thus, there was a 69% reduction in leaching time. The improved extraction data is not wholly attributable to coarse gold removal, but rather it was the combination of gold removal and rejection of other heavy mineral cyanide consumers or leach retardants. Further investigation was not warranted at this time. Preconcentration is the second manner in which leaching systems can benefit from gravity concentration. The premise is that preconcentration can reduce the quantity of leach feed, which, in turn, may reduce leaching costs. Figure 2 presents preliminary data developed by CSMRI for US Minerals Exploration (USMX). Centennial Exploration Inc., in agreement with USMX, is proceeding with evaluations to determine the suitability of various processing schemes for recovery of gold values from the Montana Tunnels property. The data shows how the ore can be preconcentrated by gravity techniques to result in a reduced feed tonnage to secondary extraction techniques, presumably flotation or cyanide leaching. Testing has shown that Reichert cones, followed by treating the cone concentrate on spirals, can deliver about 88% gold recovery in about 13% weight, that is, 87% weight rejection. Consequently, fine grinding and reagent costs are attributable to only 13% of the plant feed rate. Cost data is not yet available, but the potential exists for significant cost savings.

Jan 1, 1985

By Aloys H. Wohlrab

[THE conditions that govern the selection of a suitable type of detachable bit for the small isolated mine, for rock work and tunnel contracting and for the large mine are quite dissimilar, therefore intelligent recommendation of a particular make of detachable bit must involve consideration of many factors. Small, isolated mines may find advantageous a detachable bit that a larger mine cannot afford to use, because of more efficient conventional steel sharpening and steel-distribution practice, but a small mine may of be able to adopt the practice of larger mines because of necessary capital expendites for equipment for detachable-bit sharpening and shanking. The perfect detachable bit, suitable for all conditions, has n yet been developed. CONVENTIONAL DRILL BIT IN USE That drilling speed increases as the hit gauge decreases is well known, but since small-hole drilling is the exception rather than the rule the rate of such increase is probably not so generally recognized. Tests married out by Holman Brothers Limited1 with 7/8-in. show that when bit sizes are reduced from. 1 ½ in. 1 3/8 in., and 1 3/16 in., to 1 3/8 in., 1¼ in., and it' in., respectively, the volume of the hole is15 per cent less and the drilling speed 4q per cent higher. Recent tests by the United States Bureau of Mines2,3in hard, uniform basalt showed that a reduction in average gauge from 2.077 to 1.826 in., a difference of 0.251 in., gave an increased drilling speed of 65 per cent. The limits to which a drill hole may be extended depend upon the loss in gauge. Since the conventional bit with 5° and 14° clearances did not possess sufficient reaming edge to permit the smaller gauge changes that were required for small-hole drilling, experiments were made in bit construction with a view to correcting this deficiency. This resulted in the adoption of the full reaming cross bit (having full Carr characteristics) by some mines.4 In order to drill holes having a minimum taper from the collar to the bottom of the hole, full reaming center-hole cross bits with 90° cutting angles, 5° and I4° wing tapers, forged on 7/8-in. quarter-octagon drill rods, are used in the conventional drill-steel practice at Hollinger. In the steel shop close attention is given to the condition of the dies and dollies, which, combined with a rigid inspection of the finished bit, ensures as nearly a perfect bit as can be forged economically. Shop practices and costs have been discussed in earlier papers.5,6 The method by which the radii for the special dies used for forging bits of different gauge diameter are obtained so that the points and reaming edges cut circles of the same diameter has been discussed by Hibbert.7 REQUIREMENTS FOR DETACHABLE BIT FOR HOLLINGER Under these circumstances, it was apparent that it would be difficult to introduce a detachable bit that could successfully compete with the conventional bit already in]

Jan 1, 1942

By R. W. Luhning, S. D. Gittins, J .. A. Haston

"The Underground Test Facility (UTF) project operated by the Alberta Oil Sands Technology and Research Authority (AOSTRA) north of Fort McMurray has achieved marked success in field testing mining access and horizontal well in situ recovery of bitumen from the Athabasca oil sands.Mine access has provided for effective placing and operation of horizontal wells in the oil sands from a tunnel system in the underlying limestone. Performance of the in situ recovery process, Steam Assisted Gravity Drainage, has substantially exceeded forecasts made at the start of the project. Over-all results of the program to date show great promise for horizontal well in situ technology as a competitive source of bitumen from Alberta's vast oil sands deposits. The potential economics of a commercial expansion at the site are reviewed in this paper.IntroductionIn situ recovery o f bitumen from the Athabasca oil sands has puzzled the oil industry for many year s, but a breakthrough appears close at hand. The Underground Test Facility (UTF) project initiated by the Alberta Oil Sands Technology and Research Authority (AOSTRA) has had notable success in the piloting of horizontal well processes, which offer prospects for commercial development even under the highly competitive situation facing the Canadian industry . This paper describes the project briefly and presents the potential economics of commercial development on the UTF lease."

Jan 1, 1990

By W. L. Dennen, W. W. Wirth

Research looking toward the reduction of the cost of roof support by substitution of longer-life materials for wooden timber is fully justified by the fact that roof support is an important element of all production costs. Permanent or semipermanent roof support will reduce the expense of timber-maintenance crews during steady operations, but even more markedly during idle periods. The management, therefore, is faced with the important problem of timber renewals, which in times of regular operation are frequently obscured by more urgent problems of production. Increased use of timber in the Anthracite Region of Pennsylvania is caused mainly by increased percentage of coal obtained from robbing areas, which requires more timber for safety than development work, and maintenance of long haulageways through crushed ground despite the fact that less coal is transported over them. The timber reserve in the immediate vicinity of the mines has been depleted to such an extent that it is necessary to secure 70 per cent of the timber requirements from distant sources. Originally, mine timbering in this region was done by largely utilizing round props or legs with round collars. The whole cross section of the tree from which timber was cut was utilized by forming the joints between the legs and the collars with saw and adze, using plank or round saplings for lagging behind the legs and over the collars to prevent material falling between them. The Mine Inspectors' report of 1898 records an important digression from the use of round timber, when two tunnels at the Babylon colliery, near Pittston, Pa., were timbered with 12 by 12-in. square timber sets with mud sills, through quicksand for distances of 425 and 160 ft., respectively. The timbers were lagged with 3-in. plank to control sand. An early step toward the permanent type of roof support was the installation of "brick pillars or perfect arched brick rooms" instead of

Jan 1, 1940

By W. L. Dennen, W. W. Wirth

Research looking toward the reduction of the cost of roof support by substitution of longer-life materials for wooden timber is fully justified by the fact that roof support is an important element of all production costs. Permanent or semipermanent roof support will reduce the expense of timber-maintenance crews during steady operations, but even more markedly during idle periods. The management, therefore, is faced with the important problem of timber renewals, which in times of regular operation are frequently obscured by more urgent problems of production. Increased use of timber in the Anthracite Region of Pennsylvania is caused mainly by increased percentage of coal obtained from robbing areas, which requires more timber for safety than development work, and maintenance of long haulageways through crushed ground despite the fact that less coal is transported over them. The timber reserve in the immediate vicinity of the mines has been depleted to such an extent that it is necessary to secure 70 per cent of the timber requirements from distant sources. Originally, mine timbering in this region was done by largely utilizing round props or legs with round collars. The whole cross section of the tree from which timber was cut was utilized by forming the joints between the legs and the collars with saw and adze, using plank or round saplings for lagging behind the legs and over the collars to prevent material falling between them. The Mine Inspectors' report of 1898 records an important digression from the use of round timber, when two tunnels at the Babylon colliery, near Pittston, Pa., were timbered with 12 by 12-in. square timber sets with mud sills, through quicksand for distances of 425 and 160 ft., respectively. The timbers were lagged with 3-in. plank to control sand. An early step toward the permanent type of roof support was the installation of "brick pillars or perfect arched brick rooms" instead of

Jan 1, 1940

Discussion of the paper of F. M. Simonds and E. Z. Burns, presented at the New York Meeting, February, 1913, and printed in Bulletin No. 75, March, 1913, pp. 369 to 402. W. L. SAUNDERS, New York,. N. Y.:-Mr. Burns in his native Colorado modesty has said that he apologizes for things left out of this paper. I am much impressed by what he has put into the paper, as of all the papers that I have read on the subject, describing, as it does, a specific piece of work that has been accomplished, I do not recall one that gives so much information in such a practical way. Mr. Burns has also said that it is very hard to get men to put things down while the work is being driven. My experience is that it is very much harder to get the men in charge of work of this kind to bring them in in the form of a report before a body of engineers after the work is done. We have a great many cases where records are kept in much detail, but they are usually kept under cover, with the mistaken idea that the information gained is private information for the benefit of those who have done the work. Messrs. Simonds and Burns have opened up this case so that we can get the benefit of that experience which they have had, and so that we can learn something for the future. The description contained in this paper, as I have said, is a very practical one. You will notice in going through it. that the authors have given in detail a record of their work from the beginning to the end, and they even go so far as to publish an exact copy of the Rawley Mining Co.'s daily report. That daily report is something which those of us who are engaged in similar work might copy to great advantage, and having it before us we can check up our men while doing work and exact from them equal conditions, if it is possible to do so.

Jan 5, 1913

By Takeuchi T, Matsugaki K

We have developed a new shotcreting which is suitable for the primary support of TBM tun- nelling in soft rock. This spraying compound is made from cement, a special accelerating agent and glass fiber. It can be sprayed on to a rock surface using a specially designed gun similar to paint-spraying equipment. Some of the characteristics of this method are as follows: 1) If spraying is done keeping the distance between the gun and the rock surface be- tween 0.2 and 3.0 m, the rebound losses are small and there is little dust. 2) This method can be used on excavated rock surfaces under inflowing water conditions of up to 10 i/m2òmin. 3) After 1 day the compressive strengthes of the sprayed material is greater than 250 kgf/cm2.

Jan 1, 1987

1. What Is a Rock Burst? . . . . 569 Jack Spalding . . 569 A. F. Robertson . 569 W. R. Crane . 569 G. S. Rice......569 P. J. Shenon........... 569 L. C. Graton........... 569 2. What Are the Causes of Rock Bursts? 573 Jack Spalding..........573 A. B. Yates and P. J. Shenon 574, 578, 592, 596 A. F. Robertson . . . . . 575, 595 W. R. Crane. ... 579 G. S. Rice.........582 Oscar Weiss .........582 P. B. Bucky..........582 P. H. Bird.....588, 592, 593, 595 M. K. Hubbert.........593 G. Wheeler.......... .593 B. F. Tillson........593, 595 T. T. Read............594 3. How Can Rock Bursts Be Prevented?. 597 Jack Spalding..........597 A. B. Yates and P. J. Shenon ... 598 A. F. Robertson.......599 W. R. Crane....... .602 G. S. Rice............602 Page 4. How Can Rock Bursts Be Predicted?. 602 Jack Spalding.........602 A. F. Robertson.......602, 603 A. B. Yates and P. J. Shenon ... 602 W. R. Crane......602 G. S. Rice......602 B. F. Tillson...........603 5. How Can Rock Bursts Be Controlled? 603 Jack Spalding..........603 A. F. Robertson.......603 W. R. Crane...........604 A. B. Yates and P. J. Shenon 604 Oscar Weiss........605 G. S. Rice......- 604, 609 R.D.Hoffman..609, 610 R. D. Hall..........609 R. H. Cromwell . . 609 P. J. Shenon...........609 P. B. Bucky......609, 611 B. F. Tillson.........609, 611 A. B. Yates...610, 611 P. H. Bird.......610 6. General Discussion........ 611 Jack Spalding.......... 611 R. D. Parker....... . 616 C. Baring Horwood........ 618 P. B. Bucky........... 619 Foreword This symposium brings together points of view, experiences and ideas on rock bursts from a goodly portion of the globe. Some conceptions, particularly those regarding the necessity of depth, have been fairly well shattered, as we find that they are present in the water tunnels of New York, on the surface at marble quarries in Vermont, and in the coal mines of the United States, as well as in the deep mines of the South African and Kolar gold fields, and in the nickel, copper and gold mines of Canada. Their occurrence under such a variety of conditions is, therefore, a challenge, which, if successfully met and overcome,

Jan 1, 1946

By Philip B. Bucky

CONTENTS [PACEPACE r. What Is a Rock Burst?2 4. How Can Rock Bursts Be Predicted? . 35 Jack Spalding2Jack Spalding35 A. F. Robertson 2, 5A. F. Robertson35 W. R. Crane 2A. B. Yates and P. J. Shenon. . . . 35 G. S. Rice 2W. R. Crane 35 P. J. Shenon 2G. S. Rice 35 L. C. Graton 2B. F. Tillson 36 2. What Are the Causes of Rock Bursts? 6 5. How Can Rock Bursts Be Controlled?. 36 Jack Spalding6Jack Spalding36 A. B. Yates and P. J. Shenon . 7, 12, 29A. F. Robertson 36 A. F. Robertson 12, 28W. R. Crane37 W. R. Crane 15A. B. Yates and P. J. Shenon. . . . 37 G. S. Rice 15Oscar Weiss38 Oscar Weiss15G. S. Rice37, 42 P. B. Bucky21R. D. Hoffman42, 43 P. H. Bird 21, 26, 28R. D. Hall 42 M. K. Hubbert26R. H. Cromwell42 G. Wheeler 26. J. Shenon 42 B. F. Tillson26, 28P. B. Bucky42, 44 T. T. Read 27B. F. Tillson 42, 44 A. B. Yates 43, 44 3. How Can Rock Bursts Be Prevented? . 30P. H. Bird 43 Jack Spalding. . . . • . • • • • . 306. General Discussion . A. B. Yates and P. J. Shenon. . . . 31Jack Spalding44 A. F. Robertson32R. D. Parker49 W. R. Crane 35C. Baring Horwood 51 G. S. Rice 35P. B. Bucky 52] FOREWORD THIS symposium brings together points of view, experiences and ideas on rock bursts from a goodly portion of the globe. Some conceptions, particularly those regarding the necessity of depth, have been fairly well shattered, as we find that they are present in the water tunnels of New York, on the surface at marble quarries in Vermont, and in the coal mines of the United States, as well as in the deep mines of the South African and Kolar gold fields, and in the nickel, copper and gold mines of Canada. Their occurrence under such a variety of conditions is, therefore, a challenge,

Jan 1, 1942

By P. Soupios

Seismic geophysical methods have long been used by geologists and geophysicists to delineate subsurface features. These techniques work because different types and strengths of soil and rock transmit energy at different velocities. The Crosshole-Downhole tests (CH-DH) and seismic refraction survey (RS) are among the most used methods in engineering applications to obtain the elastic properties of subsurface layers. As a result, seismic methods have evolved into a cost-effective tool for rapidly determining depth to bedrock in engineering and construction projects. The seismic methods are best suited to sediment thickness analysis, bedrock quality determination and detection of the presence of weaknesses in the bedrock before the erection of any civil engineering structures such as bridges, tunnels, dams and portals. Seismic methods are also useful in estimating the rippability of bedrock ahead of construction. These methods are particularly useful for large-scale projects that require a significant number of drill holes, resulting in a substantial investment of time and costs for drilling. The CH-DH and RS methods are typically comparable in total cost to drilling, but provide significantly more information in 2-D and 3-D and therefore reduce the likelihood of conceptually oversimplifying the subsurface conditions. The above mentioned methodologies were considered to be the most appropriate geophysical methods for investigating the shallow structure and the dynamic soil properties of the area where the Ilarionas Dam is scheduled to be constructed. The main target of this work was to prove the potential of geophysical methods in providing accurate information to the civil engineer and to obtain information on the dynamic soil and rock properties for earthquake design analyses for the dam construction. Specifically, nine refraction seismic profiles and three cross-hole-downhole seismic tests were implemented to determine the subsurface conditions of the study area. The tests determined the shear and compression wave velocity profiles versus depth and other crucial parameters such as Young and Rigidity modules. Furthermore, they allowed the detailed 2-D mapping (along several profiles of the study area) of the subsurface variation of the soil and rock dynamic moduli.

Jan 1, 2006

By E. G. Beinlich

The continuing search for faster, safer, and more economical methods of excavating underground openings has led the staff of the Mather mine of the Negaunee Mine Co. directly to mechanical boring of raises and drifts. The mine is located within the city limits of Negaunee and Ishperning in the eastern portion of the Marquette Iron Range in the Upper Peninsula of Michigan. It is owned jointly by McLouth Steel Corp., Bethlehem Steel Corp., Republic Steel Corp., and The Cleveland-Cliffs Iron Co., who also serves as the operating agent. A total of 2,400,000 tons of iron ore averaging over 60% Fe is mined annually by the block-caving method. This provides feed for the Pioneer pellet plant at Eagle Mills where 1,450,000 tons of pellets are produced yearly with the remaining tonnage being sold as coarse ore. Technical staff of Cleveland-Cliffs have long realized that improved methods of advancing raises and drifts were a necessity if underground mining of iron ore was to survive the transition period in the steel industry's requirements from direct-shipping ore to pelletized ores. For this reason, the Mather mine embarked upon an experimental raise-boring program with a prototype machine in 1961. Improvements since that time in equipment and practice have resulted in the acceptance of mechanical boring of raises as a standard development tool and a vital link in the fulfillment of an orderly development schedule. Advancement of large-diameter openings such as main-level drift headings by mechanical methods in the harder, more competent types of rock has developed more slowly in the mining industry due to the need for more compact and portable machinery and improved bit technology. Many of the obstacles have been overcome by the manufacturers, and in March 1967 it was decided to proceed with a program to bore approximately 4000 ft of 13-ft-diam drift in the Mather graywacke footwall rock.

Jan 1, 1970

The four major contracts for the construction of Museum Station are: 1) Installation of Soldier Piles, 2) Structural Contract including the temporary support, 3) Architectural Finishes which include Electrical and Mechanical Services and 4) Supply and Installation of lifts and escalators. Specific areas of interest include the fabrication, installation and removal of temporary support steelwork, the monitoring of ground loading, and the interface between the structure and. the temporary support. The information collated during construction confirmed the design assumptions and criteria. No major changes'were required in the design of the construction methods. The open cut excavation and construction of this major structure in an inner city area was significant by any standard and was successfully completed with minimum disruption to the public and without damage to adjoining or surrounding property.

Jan 1, 1981