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By C. L. Barron, M. L. Smith, K. A. Prisbrey

SAG mill throughput is often a bottleneck to increasing plant productivity. Improved size distribution control in the mill feed can be used to increase mill production. This can be accomplished by controlling blast fragmentation as a function of energy input and rock characteristics. Such an approach requires the integration of rock fragmentation and comminution modeling. The optimal mill feed size distribution is determined by adapting a population balance comminution model to a measure of rock hardness or fragility, which is also related to the rock fragmentation that is achieved by blasting. The relationship between the size distribution and fragility of the feed and the mill's product and the critical size in a SAG mill is examined.

Jan 1, 1994

By Robert H. Merrill

INTRODUCTION In most cases, the design of an underground mine is based upon the premise that the ground either will cave or will be stable. This chapter concerns the design of a mine in ground that will cave readily or with some as¬sistance, such as by long-hole drilling and blasting. Some of the more widely used caving systems of mining are panel caving, block caving, sublevel caving, and large pillar recovery. Some of the less widely used systems are glory-hole, top slicing, and induction caving. Al¬though the common practice of pillar robbing is not usually considered to be a caving system, this subject will be treated as a part of this chapter. BASICS OF CAVING Caving systems are most successful in ground that will cave in sizes that will flow through openings and grizzlies, and will easily load in cars or on belts for haul¬age. The ground most likely to cave well is highly frac¬tured and contains breaks, flaws, or other discontinui¬ties that form planes of weakness. Also, caving action can be greatly enhanced if the host rock itself is low in compressive, shear, and tensile strength. Ideally, a cav¬ing system of mining is best employed when the criteria for caving is a feature of the ore body and the develop¬ment drifts, haulageways, and drawpoints can be mined in a highly competent rock beneath the mineralized zone. However, the development is often in the same, or similar, fractured rock and the openings require sub¬stantial artificial support to assure stability. Several clues can be assembled to identify potential caving ground; however, for borderline cases, no sure method has been devised to date. The diamond-drill cores taken for exploration can provide an excellent clue provided drilling is performed carefully by experienced drillers. For example, if the ground is cored in such a manner that the breaks in the core are caused more by failure of the rock than by whipping core barrels, plugged drill bits, or other drilling causes, and the intact core lengths are consistently long [say, 0.6 to 3 m (2 to 10 ft) of unbroken core], there is little reason to believe the ground will cave without considerable as¬sistance. This is especially true for rocks with compres¬sive strengths above 34.5 MPa (5000 psi) and tensile strengths above 2.1 MPa (300 psi). On the other hand, if core recovery is low (below 80%) and the recovered ore is broken in small pieces and the breaks are along obvious weaknesses in the rock, the chances are excel¬lent that the ground will cave. This is true even when the rock between the defects has high compressive and tensile strength. Another clue has already been mentioned, that is, the measurement of the physical properties of the rock and the natural planes of weakness or defects in the rock. The planes of weakness in the rock can often be detected from outcrops, cores, or other exposures of the rock under consideration. Some rock types are known to be strong and will sustain large, unsupported open¬ings and would be difficult to cave intentionally. Yet the same rock type can also contain unbonded or weak planes of weakness or fractures, and in these locations the rock would undoubtedly cave with little assistance. Therefore, although the inherent strength of the rock is a factor in caving, the natural defects in the rock are more often the deciding factor. DESIGN CONCEPTS For the most part, the design of openings for caving ground is a problem of the interaction of openings over a relatively large area of the mine. To illustrate, Fig. 1 is a simplified section of a series of openings along the grizzly level or draw level of a block caving or panel caving development, and above this opening is a simpli¬fied section of a room-and-pillar arrangement on the undercut level. At this stage of the development, the stresses around the openings on the grizzly level are only moderately influenced by the openings on the undercut level and vice versa. Therefore, the stresses around the openings are approximated by the stresses around single or multiple openings in rock, the values of which are de¬scribed in the literature (Obert, Duvall, and Merrill, 1960; Obert and Duvall, 1967). Once the pillars on the undercut level are blasted (Fig. 2), the situation changes abruptly. The undercut opening (prior to caving) now can be approximated as an ovaloidal opening above the grizzly drifts and this opening tends to shield the vertical stress field. As the caved stage is drawn the stope approximates a much larger rectangular or square opening filled with rock, and if the rock is not sustaining a major portion of the stress field, this opening can be considered (for en¬gineering purposes) to be empty and the stresses that interact between the larger and the smaller openings take on a totally new perspective (see Fig. 3). Next, let the material cave to the surface, and let the caving ma¬terial sustain some stress, but much less than if the ma¬terial were intact. This condition is similar to a soft inclusion in a rigid body and has been treated in the literature (for example, Donnell, 1941). At this point in time, the grizzly drifts are subjected to the stress con-

Jan 1, 1982

By Wisecarver D. W.

Today the longwall is getting increasing attention in the United States as an excellent way to mine coal. This is not to say it is the best way in every mine because conditions, capital costs, and markets may well dictate another method. However, operators with working longwalls are often showing preference by phasing out room-and-pillar sections and replacing them with longwall. In the United States there are about 90 longwalls in operating coal mines and the number is increasing. In Europe, 95% of underground coal mining is by longwall compared with only 5% in the United States. In the Western States about 25% of the underground coal mines use longwall mining. There is logic behind the increase in longwalling. Compared to room-and-pillar mining, straight single-pass longwall has important advantages that show up in underground coal mining in any part of the country.

Jan 1, 1978

By Steve Kral

The Red Dog zinc-lead deposit, said to be the world's second largest, is located inside the Arctic Circle in northwestern Alaska - a location too harsh and remote for some mining companies but just right for Cominco Ltd. The complex has been in operation since November 1989 and is producing 508 kt/a (560,000 stpy) of zinc concentrate, 109 kt/a (120,000 stpy) of leadconcentrate and 45 kt/a (50,000 stpy) of an Imperial Smelter Feed, a combined zinc-lead mixture that is too small to separate. All of this concentrate production derives from a relatively modest mine production of 5.4 kt/ad (6000 stpd) extracted from a surface pit. Despite being a high-grade zinc-lead deposit, Red Dog's metallurgy is complex, requiring a very fine grind and use of column cells for flotation. The Red Dog complex is located in the DeLong Mountains. They are characterized by eight stack and folded thrust blocks that contain sedimentary and igneous rocks. The Red Dog deposit is located in the second lowest thrust and ist hosted by late Mississippian to Permian age black shale, the Ikalukrok member of the Kuna formation. The Red Dog mining complex is operated by Cominco Alaska, Inc., a wholly owned subsidiary of Cominco American, Inc, which, in turn, is owned by Cominco Ltd. In 1980, the deposit region was acquired by NANA Corp. as part of the Native American Claims Act Settlement of 1971. NANA is one of 12 such Native Corporations in Alaska. In 1982, NANA and Cominco American signed a develop¬ment agreement and formed Cominco Alaska to manage the project.

Jan 1, 1992

By L. M. English, Y. J. Wang

When plotted as head vs. volume, fan curves start high and decrease with increasing volume, while mine head curves start low and rise with increasing volume. Their intersection is the system operating point. However, as every ventilation instructor knows, the fact that head loss is a negative number (debit) and fan head is an energy addition (credit) is often a stumbling block to student or audience comprehension since the curves seem to reflect the opposite. It is possible to use the intuitively correct curves- pressure losses as negative values and energy additions as positive values - and still obtain the correct operating point. Adding the two curves produces a third curve that intersects the x-axis at the operating point, the point at which the fan energy input exactly matches air circulation energy loss. Using this technique also makes it easier to account for outside influences on the system, such as natural ventilating pressure. With the standard approach, it is quite easy to add or subtract such influences to the wrong curve. This approach can be extended to the graphical solution of any series/parallel ventilation network. The results are the same, and the approach is logically consistent and less confusing.

Jan 1, 1999

By William J. McCloy, Robert L. Gamble

A coal-fired fluidized bed steam generator rated at 12.6 kg/s has demonstrated the industrial and institutional application of fluidized bed combustion by burning high sulfur coal in an environmentally acceptable manner within a populated area. Designed to provide steam for heating and cooling, the unit was placed into service at Georgetown University in Washington, DC, and has already logged 5275 hours.* The unit has consistently operated well within Washington's strict regulations governing emissions of sulfur dioxide (SO2), nitrogen oxides (NOx), and particulates. The Georgetown program, cofunded by the university and the US Department of Energy, has shown that fluidized bed combustion is a promising way to realize the full potential of coal, the nation's most abundant energy resource. In a fluidized bed steam generator, a layer of granular limestone is placed on top of a grid designed to prevent particles from falling through. Air is passed upward through the bed with sufficient force to lift and float the particles. At this point the bed material resembles a boiling liquid. Crushed coal is then introduced into the bed and ignited. When high sulfur coal is burned, limestone absorbs the SO2 produced. The operating temperature of a fluidized bed steam generator is relatively low (approximately 871°C). As a result, NO, formation is minimal. Another potential problem encountered in burning coal is fly ash, a form of particulate. A feature of the fluidized bed design, developed by Foster Wheeler Energy Corp., is a fly ash reinjection system. Fly ash is reintroduced into the bed where it is burned again, thereby enhancing fuel efficiency. Nearly all remaining fly ash is captured by a baghouse dust collector. At Georgetown, the fluidized bed steam generator and all auxiliary equipment are enclosed in a structure aesthetically compatible with the surrounding university complex. The steam generator has two separate beds that can be fired independently, and operation to date has been with one and two beds in service, based on steam load demand. During several operating periods, when the fluidized bed unit was the only boiler in service, it supplied all the steam required by the 18.5 hm2 of buildings on campus. Initially fired with low sulfur coal, the unit was quickly switched over to higher sulfur fuel which will be used most of the time.

Jan 3, 1981

By Gregory E. Klosterman

American Aggregates Corporation is one of the leading producers of mineral aggregates in the nation. Production is from deposits of glacial sand and gravel and limestone and dolomite bedrock. Annual production from 33 plants is 25,000,000 tons. Land consumption is at the rate of 300 to 400 acres annually. This annual rate of consumption presents a continuing challenge to the exploration and mining personnel to meet the requirements of reserve replacement, mining and land rehabilitation.

Jan 1, 1974

By DeVerle P. Harris

The structure of an exploration model reflects a number of design considerations, one of which is the question or issue-to be investigated by the model. Thus, a model designed to evaluate the magnitude of risk and its economic implication to optimal decision making at each decision juncture of a multistage exploration architecture differs considerably from one designed to evaluate, say, the long run average cost of discovery, in which the architecture of exploration is suppressed and risk is simply an element of average cost. The perspective of the model examined in this paper is that exploration is investment in (purchase of) information and that the next stage of an exploration program is undertaken only when the expected value of information to be acquired in that stage exceeds its cost. The degree to which this perspective is honored influences considerably model design. Consider two design objectives:

Jan 1, 1989

By C. Huang, S. Lin, Y. J. Wang

Introduction A mine characteristic curve is of great importance for both theoretical analysis and daily management of a mine ventilation network. Ventilation engineers are well acquainted with the ways of obtaining and using a mine characteristic curve that is conventionally associated only with single-fan networks. However, for a multiple-fan network, difficulty has long been recognized in defining the curve, not to mention, obtaining or using it. To extend the concept of mine characteristic curves for a multiple-fan network, a subsystem-based approach has been introduced by Wang (1992a). He defines a subsystem characteristic curve for a multiple-fan network and discusses the numerical procedure for constructing the curve. A graphical method of obtaining a subsystem characteristic curve for serial-parallel, multiple-fan networks has also been accomplished (Wang, 1992b). In spite of the maturity of the fundamental concepts related to and the numerical analysis tool for a multiple-fan ventilation network, information regarding the related laboratory tests is very scarce. To make up this gap, laboratory experiments have been done by using a two-fan experimental network. This paper presents the methodology and major findings regarding the construction and features of a subsystem characteristic curve. Characteristic curves and measurement fundamentals A characteristic curve A characteristic curve is a graphic presentation of the relationship between any two variables of a two-terminal component. It is typically defined in the form of a quantity-pressure curve for a ventilation component, which can be a single branch (airway), a fan, a regulator, a portion of a network or even a whole network. If a component is a single branch, the curve is called a branch characteristic curve. If it is a fan, then the curve is called a fan characteristic curve, and so on. Figure I illustrates the characteristic curves of some typical ventilation components. A general method for obtaining a characteristic curve from measurement is to apply an adjustable-pressure source between the two terminals of a component and measure the quantities corresponding to the different pressures. Mine characteristic curve A single-fan network can be easily divided into two ventilation components, the fan and the remaining network. The latter consists of only branches and regulators. Under this subdivision scheme, two characteristic curves can be constructed. They are the fan characteristic curve and the characteristic curve of the remaining network. The latter is referred to as the mine characteristic curve or the system characteristic curve. Although the general method for constructing a characteristic curve from measurement can be applied to obtain a mine characteristic curve, there is a shortcut for constructing the curve by introducing a mine resistance factor R. It is a constant if all of the resistance factors in the network are constant. R relates mine pressure H with quantity Q as in: H=RQ2 (1) For a given R, Eq. (1) can be directly used to plot a mine characteristic curve by assigning a series of Q values and calculating the corresponding H values. Hence a shortcut to constructing a mine characteristic curve is to measure one set of mine pressure H and quantity Q values, calculate the mine resistance factor R using Eq. (1) and then plot the curve. Subsystem characteristic curve Applying a similar scheme as is used for a single-fan system, divide a multiple-fan network into two components, a fan under consideration and the remaining network that contains all the other elements (branches, regulators and remaining fans) of the network. The fan under consideration

Jan 1, 1994

In the mining industry and in new and expanding fields, such as electronics, nucleonics, and astronautics, a new type of manager has evolved. The name most generally used is project manager. His business is to create a product and the principal means through which he accomplishes this is project management. In this chapter, new information is the product of the project manager; as such the products, for which the computer methods described throughout this volume are designed, fall in the domain of project management. It has been the author's experience that project management is the management tool to effectively direct the information technology function within an organization. The purpose of this chapter is to describe project management and to identify and present a solution to the problems associated with project management in the development and application of computer methods in a minerals environment. Specifically, the problems to be addressed will center around: (1) recognition of the need for project management in terms of business-environmental factors and product characteristics, (2) formulation of a method of operation in terms of organization and implementation, and (3) application of project management to the minerals industry.

Jan 1, 1979

By Carl Eschman

Productivity from large walking draglines is primarily dependent on operator skills. This machine may be in operation three shifts a day, 364 days a year, and its output is directly related to coal uncovered and mine profitability. Dragline operators must have highly developed manual skills and be knowledgeable in mine planning and working strategies. When using equipment costing more than $25 million, some formal training is usually required before an operator is allowed to assume complete con¬trol; however, dragline operators rarely receive any structured training before operating these giant excavators. A form of apprenticeship is usually followed where an operator candidate progresses from a groundsman to an oiler position. As an oiler, he is permitted to operate the dragline for short periods under supervision. After apprenticeship, the operator is considered sufficiently prepared to operate the largest, most powerful machine at the mine. The apprenticeship training method has obviously provided the surface mining industry with skilled dragline operators; however, conditions are arising that require a realistic and effective training tool that can be accessed by mining companies. New mines -either planned, under construction, or recently opened in the West-do not have access to a pool of experienced operators and oilers as do Midwest mines. As coal mining activities increase in both the West and Midwest, demand for trained dragline operators could be required in a short amount of time. Also, the more productive techniques along with sound basics of strip mining are sometimes lost in the informal "OJT" training method. Modern draglines are the pacemakers of the strip mine, and are simply too expensive to be used as a training device where lost productivity and susceptibility to damage can directly affect mine output. The Dragline Training System is a logical first step in formally educating or retraining operators. The program, started by the US Bureau of Mines and continued by McDonnell Douglas Electronics Co. under contract with the US Department of Energy, was installed and evaluated at DOE's Carbondale Mining Technology Center in Carterville, IL, last year. It is now being operated by Southern Illinois University at Carbondale. System Description The Dragline Training System addresses specific environments and work practices encountered during an actual mining operation at a midwestern US surface mine. This area was chosen because of its high number of strip mines using large walking draglines. Most draglines in the region are Bucyrus-Erie 1370, 1450, and 2570 models, so the dragline trainer was patterned after the company's 1370 machine. Operating and emergency controls are sufficiently standardized for most large walking draglines, and peculiar dynamics and responses from any specific dragline can be programmed into the computer system. The computer simply prompts the user to select the manufacturer and any peculiar response or rate changes needed. A 46-m3 (60-cu yd) bucket is simulated, but for closer simulation, various bucket configurations can be provided. Dragline Trainer The dragline trainer uses the TV-model simulation technique. A scaled model of the dragline is positioned in a model mine. A television camera is positioned at the operator's theoretical eyepoint, and the view captured is projected into a large screen in full scale. The screen is positioned in front of the operator seated in a full-size cab at Bucyrus-Erie controls. By manipulating the controls, the trainee can operate the model dragline and observe its reaction in the television display. In addition to housing dragline controls and consoles, the wooden, oversized cab contains the digital computer, terminal, video recorder and monitor, power switch box, air conditioner, and has enough room for the instructor and five student observers. The 50:1 scaled dragline model contains servo-controlled functions for hoist, drag, swing, delta swing, and longitudinal and lateral position. The delta swing provides bucket lag during swing and a realistic pendulum action when the swing is terminated. The over-responsive second order servo system is designed to provide hoist, drag, and swing rates exceeding present draglines. In all cases, position servos are used for better control and sta¬bility. The normal rate operation of an actual dragline is computed for the specific machine and presented to the servo amplifiers as iterative position commands increasing or decreasing

Jan 6, 1982

By Dale R. Ralston, Roy E. Williams, Gerry V. Winter, George L. Bloomsburg

INTRODUCTION This manuscript is based on site inspections of five existing mines and the detailed analysis of two of the five. The detailed analysis and mathematical models of the two mines are presented herein in order to demonstrate a method of predicting mine water inflow using existing mines as a data base. Water management procedures were studied at all mines. Our original criteria for mine selection were that the mines be less than five years old and have an inflow rate in excess of 2000 gpm (7.6 m3/min). Background hydrologic and hydrogeologic data were required so that two of the mines could be used for our computer modeling experiments. Our original criteria proved to be difficult to meet. We contacted numerous mines via various sources. Many mines were unable to meet the cri¬teria stated or were not interested in cooperating with us. Therefore, we allowed some flexibility in the application of these criteria to satisfy our orig¬inal requirements. Although the visitations at all the mines are not discussed herein the characteristics of the two mines simulated are presented. These simulated mines consist of a uranium mine in New Mexico and a lead mine in Missouri. The mines and owners will remain anonymous by request. URANIUM MINE-CASE HISTORY NUMBER 1 Mine Location The uranium mine that we simulated is located in northwestern New Mexico. The ground surface at the uranium mine is at an elevation of approx¬imately 7000 ft (2134 m) above mean sea level. The climate is arid to semi-arid with an average annual precipitation of approximately 10 in. (25.4 cm) (John and West, 1963). Geologic Setting It is essential that a conceptual model of the stratigraphy, hydrostratigraphy, and structural geology of a mine site be developed prior to simulation by a mathematical model. If such a conceptual model is not developed prior to simulation major errors are difficult to avoid. The uranium mine we have simulated is located in the San Juan Basin (Fig. 68), which is a structural basin located in northwestern New Mexico and adjacent parts of Colorado, Arizona, and Utah. The sedimentary rocks present in the deepest part of the basin are nearly 15,000 ft (4572 m) thick (John and West, 1963). The regional dip of the sedimen¬tary rocks in the Grants area is generally northward toward the San Juan Basin; the regional attitude is modified locally by normal faults and folds (Hilbert, 1963). The rocks of interest in this area are sedimentary in origin and range in age from Penn¬sylvanian to Cretaceous. They rest on the Precam¬brian core of the Zuni Uplift. Associated intrusive and extrusive rocks are of Tertiary and Quaternary ages (Hilbert, 1963). The deepest sedimentary unit of interest at this mine site is the Morrison Formation as described by Hilbert (1963). The Morrison Formation con¬sists of three members in the general area of the mine site; the deepest member is the Recapture. The Recapture Member consists of alternating beds of gray sandstone and grayish red siltstone or mud¬stone with strata up to several feet in thickness. Only a few small deposits of uranium are contained within the Recapture Member. The Westwater Canyon Member overlies the Recapture Member of the Morrison Formation. The Westwater Can¬yon Member is approximately 150 ft (46 m) thick and consists of light brown to gray, poorly sorted, arkosic sandstone, and some interbedded gray mud¬stone. The Poison Canyon Formation denoted in subsequent discussions, tables, and figures was de¬termined to be a part of the Westwater Canyon Member. The separate connotation is used for the description of mining operations and hydraulic properties but a separate physical description is not included. This member of the Morrison Formation contains many large deposits of uranium. This is (some pages in this article are damaged)

Jan 1, 1986

By L. Vandamme, J. H. Curran

A fully three-dimensional hydraulic fracture propagation model, based on the Displacement Discontinuity Technique, is briefly described. Using this model, it is shown that the geometry of a propagating fracture is dependent on several factors, among which the fluid pressure distribution within the fracture plays an important role. The containment of hydraulic fractures is found to be dependent on stress and toughness barriers in the formation. Containment is also observed to be more effective with a decrease of the viscosity, pressure or flow rate of the injected fluid. The three-dimensional layout of the fracture is shown to have a definite influence on the width and pressure distributions within the fracture. The examples considered indicate the need for fully three-dimensional models, as complements to the commonly used two-dimensional and pseudo-three-dimensional models.

Jan 1, 1986

By L. E. Baker, M. J. Leck

OSIRIS is a continuous monitoring system for respirable dust which has been specifically designed for mining and related industries. By using the latest microprocessor technology, OSIRIS provides a sophisticated dust measuring system which is a cost effective alternative to regular gravimetric sampling. The system is certified intrinsically safe for use underground. Up to eight underground OSIRIS sensors can be controlled and monitored by a single IBM PC computer located at the mine surface. This computer provides continuous and instantaneous information about the dust concentrations at the sites being monitored. Information is available in tabular and graphical form. The paper gives details of proving trials of the OSIRIS system in some UK coal mines. During these trials, adjacent OSIRIS sensors placed on the same underground roadway cross-section gave virtually identical results. Throughout a good correlation was obtained with conventional gravimetric dust samples. The OSIRIS equipment proved both reliable and easy to operate. OSIRIS was designed and developed by the Safety Engineering Laboratory of the UK Health and Safety Executive. It is now available commercially and is being

Jan 1, 1987

By Feng Cai

"As a new method to support gate roadways, rock bolting has been widely used in the big coal enterprises for its superiority. The disturbed stress field of surrounding rock and anchor pre-stress field were divided into 3 areas in consideration of s1, s3. The surrounding rock failure form and characteristic were analyzed in combination with parameter f, the failure mechanism was explained using the slip line theory, furthermore, the rock bolting mechanism and its arrangement were studied. After roadway excavation, the tensile stress area appeared at the surface of the surrounding rock. The failure type changes from shallow tension-shear form into deep compression-shear form and the failure shape is an “incense burner”, surrounding by high risk failure area in butterfly shape. The maximum principle stress in roof appears discontinuous distribution because of layers separation. The failure type of surrounding rock is intensity-stress environment-weak plane control model. The internal reason of roadway macroscopic failure is the displacement between the surface plastic area and the deep elastic region along the tangential direction of slip line. In condition of rock bolt support, there is one-way tensile stress area in small range at the surface of surrounding rock, inside is the bidirectional compressive zone in uniform distribution form, stress environment is improved by the superposition of anchor pre-stress and disturbed stress field, anchor rock bolt arrangement and pre-tightening force distribution is decided according to the distribution of f parameter. Finally, supporting method and parameter reasonability was proved by the roadway deformation monitoring, damage depth drilling peep test in the 8101 transportation roadway in a mine.Introduction The stability of roadway support is very important to the safety and sustainable development of coal mine. In recent years, the focus of China's coal mining industry began to move to West China, the thick-extra thick coal seam is developed in the West. The latest technology and equipment have been put in to service, the occurrence condition of thick coal seam and the increasing of productivity lead to the cross section of the roadway increasing continually, hence, it is more difficult to control the surrounding rock. At present, the most common way to support the roadway is rock bolt (cable) support (Wang 2007); Scholars have been deeply studied on the mechanism of rock bolt. It is reported (Wang et al 2012, Gao 2007, Zhang and Yuan 2006) that the influence of rock bolt on physical and mechanical parameters of coal and rock mass in the anchorage zone, It is believed that the rock bolt can improve the peak strength of surrounding rock, especially the post peak strength, and the failure mode of rock mass under the rock bolt support is analyzed. It is reported (Wang et al 2014, Xiao et al 2013, Li et al 2007, Yan et al 2012), according to different geological conditions, the failure mechanism of surrounding rock and supporting technology of the whole coal roadway were studied. Focus on the failure mechanism of roadway surrounding rock, different forms of high strength support methods were proposed, which achieved good results. It is reported (Wang et al 2012a, Zhang et al 2011, Wang et al 2012b, Kang et al 2010, Zhang et al 2010, Kang et al 2007) that the supporting mechanism of pre-stressed anchor was analyzed from the point of the distribution of the maximum principal stress in the pre-stressed anchor field, and the influence of the rock bolt support parameters and surrounding rock characteristics on the supporting effects were studied. Based on the latest research results, this paper analyzed the failure mechanism of surrounding rock of coal roadway, proposed the high strength anchor net cable combined supporting technology based on the feature of stress zoning distribution of surrounding rock, to provide reference for the coal roadway rock bolt (cable) support technology."

Jan 1, 2017

By D. W. Fuerstenau, G. S. Sudhir, P. C. Kapur

Quartz-limestone feed mixtures were ground in a laboratory-size, choke-fed, high- pressure roll mill. As compared to individually ground minerals, the harder component (quartz) in the product mixture is coarser, and the softer component (limestone) is finer. An energy split index is introduced for quantifying the disproportionate split of the mill energy between the feed constituents. This index is incorporated into a grinding kinetic model for simulation of the product size distributions of the mineral components as a function of energy input to the mill. The self-similar character of the size distributions is analyzed and an energy-size reduction relationship for the roll mill is highlighted.

Jan 1, 1992

By I. Nishkov, L. Alexandrova

The kinetics of froth flotation and the three-phase contact (TPC) parameters are studied at temperatures varying from 20°C to 40°C. The material floated consists of glass spheres and a quartz particles with various diameters. Dodecylamine hydrocloride (DDAHCI) is used as a collector. The equilibrium three-phase contact angle is measured with a microscope technique using the modified Mingins-Scheludko method. Using a Wilhelmy method, the temperature effect on the surface tension of the DDAHCI solution is determined. The formation of the thin film on a polished plate of vitreous silica and the expansion of TPC is monitored by means of a metallographic microscope and recorded on a attached camera at different temperatures. The experimental data show an intensification of the flotation kinetics at 30 and 40°C in the case of small particles. Such an intensification is not observed with the larger particles. The cause for intensification of the flotation kinetics is sought for in the dependence of the TPC rate of expansion and induction time on temperature.

Jan 1, 1995

By Donald J. McKee

The product sizing and tonnage from a grinding circuit are the end result of a series of linked size reduction stages which always involve: fragmentation in mining; degradation in transport of ore; crushing and finally grinding. A case study is presented which illustrates the substantial effect on autogenous mill throughput of different mining practice. It is shown that it is possible to model each of the size reduction steps of fragmentation, degradation, crushing and grinding. The way in which this predictive capability is used is dependent on the operating constraints of each mine-mill combination.

Jan 1, 1998

By K. Bragg

During the last several years considerable discussion has taken place on the long term aspects of uranium tailings management. These discussions have centered around a variety of topics which in turn lead to a variety of opinions. In order to come to grips with the various differences of opinion and points of view, the Atomic Energy Control Board has decided to prepare, as an interim measure, a list of proposed close-out criteria for uranium tailings management facilities. It is important to keep in mind that the criteria set out in this paper are proposals and are intended to initiate a continuing dialogue between the uranium mining industry and the various regulatory and government agencies. It is envisaged that this dialogue would continue for a period of from six to twelve months after which some final form of these criteria would be issued and applied to the active uranium mining facilities. In attempting to understand the reasons why the Atomic Energy Control Board has decided to propose these criteria now instead of at some future point in time, it is necessary to look at a number of existing facts. One of the most important of these is that something in excess of one hundred million tonnes of tailings currently are placed in various tailings management facilities. Furthermore, the rate at which tailings are being deposited into such management facilities is on the increase and this increase is expected to continue over the next five to ten years. The management of these tailings is currently covered under the Mine Facility Operating Licences issued by the AECB. Implied within these licences and supporting documents leading to them, are requirements for the conceptual approach to be used at abandonment. There is however, no current requirement for a detailed study on the engineering or financial aspects of such abandonment procedures. This is a gap which the Board feels must be addressed. There are, at this time, various initiatives underway both in Canada and within the international community which are attempting to address the problems and concerns of the very long term management of tailings. For example, current research in Canada which is composed of two components, one by the AECB dealing with purely regulatory aspects and a second major coordinated program under the direction of the Department of Energy, Mines and Resources and in particular by CANMET is going to deal with the technical problems. On the international scene

Jan 1, 1980

By Gerrit V. R. Goodman, Fred N. Kissell

The issue of miner safety in case of an underground mine fire has been extensively examined by both government and industry. These efforts have intensified in recent years following a number of fatal fires at underground coal mines. As a result, there is an increasing interest in the reliability of coal mine escapeways during a mine fire. This paper discusses the use of fault tree analysis for evaluating escapeway reliability. A fault tree is developed using standard logic formulations. It is later evaluated using commercially available software. The computer model uses Boolean reduction to obtain the minimal cut sets and minimal path sets of the tree. A brief analysis is also conducted to determine the occurrence probability of the TOP event. This study shows that fault tree analysis is a viable means of analyzing escapeway reliability. The cut and path sets reveal those event sequences leading to fatalities. Such information is useful in safety training and accident prevention. The probability of the TOP event shows the frequency of occurrence for fatalities in an underground mine fire. The usefulness of this value, however, is limited by the accuracy of the data for the initiating events. For mining events, the data is difficult to estimate precisely.

Jan 1, 1989