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By Duncan M. Lamb

Commercial computer "time-sharing" is a new and powerful tool for the engineer. It is a system of virtually simultaneous use of a digital computer by several individuals. Access to the computer is by means of a conventional teletypewriter which transmits over the regular telephone lines to the computer which may be located any distance from the user. Complex or repetitive computations can be programmed and the speed of computation is many times that which one can do with slide rule and desk calculator. The availability of such a commercial computer "time-sharing" system offered an opportunity to transform a tedious and time-consuming task into one that requires only a few minutes. Until the work of F. C. Bond produced widely accepted mathematical statements concerning the performance of grinding mills, the selection of the proper size equipment was done by comparison with similar operating equipment. Allis-Chalmers systematized Bond's work in a bulletin published in 1963 but the calculations are laborious and the final mill selection still required comparison of the calculated power draw with the power requirements listed in equipment catalogs. This was, however, a step forward from previous procedures. This paper discusses a rapid and simple method of determining the size and power requirement of overflow ball mills. A shared-time computer program has been developed which rapidly calculates and displays the physical dimensions, power requirements, grinding media size and quantity, as well as the important correction factors used. All of these data are calculated from input information which is usually available.

Jan 1, 1969

By St. Mergentime, K. Zeidler, V. Gall

2001, the New Jersey Transit Authority commissioned the JV Merco-Obayashi with the rehabilitation contract of the approximately 130-year-old, 4,200 ft long, partially brick- lined Bergen North Tunnel. The contract entailed the tunnel enlargement, shotcrete and rock dowels initial support according to the NATM, a waterproofing system, tunnel drainage and a cast-in-place concrete lining. The rock conditions encountered required design adaptations and significant modifications of the initial tunnel support. Flexibility, skill and extensive experience on the Authority’s and Contractor’s side finally lead to a success. This paper describes the operations and design modifications carried out.

Jan 1, 2003

By J. Gonzales

For this paper, blending is conceived as a linear process that takes a proportion from one stockpile or location and mixes it with another proportion from another stockpile or location, with the aim of obtaining a set of attributes in the final combination that satisfies a given quality standard. A blending class is the combination of ore attributes and grade limits that should be stockpiled in order to achieve the target quality for the process Plant. Current literature considers the number of blending classes in the Open Pit Scheduling Problem as fixed; however, this paper shows how to define specific blending classes for a given deposit so that the schedule can be solved with Sequential Linear Programming. Finally, a practical application in a Peruvian copper mine is shown, involving four ore attributes: alteration code, arsenic, oxide content and copper grade. INTRODUCTION In copper ore deposits, blending for quality has not been a typical issue as it was for coal and iron ore; however, current copper deposits in the world are becoming geologically and chemically more complex than in the past and is not uncommon to have quality standards involving up to four attributes, introducing an additional complexity to the mine scheduling process [1]. In order to solve the Open Pit Scheduling Problem with multiple ore quality constraints, two assumptions are necessary [2]: - Grade constraints are defined only by specifying an upper and lower limit; - Ore attributes in stockpiles are pre-determined, i.e., the upper and/or lower grade limits defining each stockpile are already known at the time of setting the schedule. Assuming each ore attribute is classified as high, medium or low, and there are four attributes to be controlled in the target ore quality, the number of possible blending classes would be 3^4 = 81. This number is impractical to manage in Operations, and increases the complexity at the time of running the schedule; therefore, a more efficient way of classifying ore is required. Note that in practice, one blending class can be split into more than one stockpile location due to physical capacity constraints. Therefore, the number of stockpiles could even be more than 81. BLENDING CLASSES DETERMINATION For a given set of ore grades with lower and upper limits to be achieved, one has to find all feasible ore classes such that when blended, their grades will lie within those limits.

Jan 1, 2019

By Pradip

The development of highly selective mineral processing reagents, which are not merely metal-ion specific, but specific to the particular mineral surface structure, is essential for the exploitation of relatively more difficult-to-process ore deposits, in particular for the separation amongst minerals having the same constituent ion. Both the choice of the functional group as well as the molecular architecture of the reagent determine its selectivity. Several examples are presented in this paper to illustrate the concepts of structure specificity and molecular recognition at mineral surfaces - the concepts crucial to the advancement of the emerging science of reagents design. Recent advances made in the understanding of molecular interactions involved in the synthesis of biomineralized materials are discussed so as to delineate the utility of this approach in the design of selective reagents for mineral separation.

Jan 1, 1994

By G. W. Riley, D. E. Pickett

HINDERED SETTLING CONCENTRATION In the free settling of mineral particles in a liquid, the falling particles are at a distance from each other so that no particle is affected by its neighbor. In hindered settling, the concentration of particles is sufficiently high so that each particle is affected by its proximity to other particles in the suspension. Richards and Locke86 have described the hindered settling phenomenon as the condition “.. where particles of mixed sizes, shapes and densities in a crowded mass, yet free to move along themselves, are sorted in a rising current of water, the velocity of which is much less than the free-falling velocity of the particles but yet fast enough so the particles are in motion." This is the condition normally encountered in mineral con¬centration processes. The well known Newton equation for free settling of coarse (ap proximately +10-mesh, 1/16-in. or -- 2000-µm) spherical particles is:87 where v," is free settling velocity, cm/sec; p' is density of the fluid; p is density of the particle, g/cm3; d is particle diameter, cm; g is acceleration due to gravity, cm/sec2; and Q is coefficient of resistance, dimensionless, ~0.4. For the settling of fine spheres in water (approximately 150 mesh or 100 µm) the equation of Stokes pertains:87 where µ is viscosity of the fluid in poises and the other symbols have the same meaning as those in Eq. 1. For particles whose size lies between about 10 mesh (--2000 µm) and 150 mesh (100 µm), their settling velocity can be determined from experimental data. These data are available in convenient form in the text by Taggart88 based on the original work of Richards. Alternatively, a Reynolds number-coefficient of resistance plot may be used to determine the settling rate of such particles.87 The settling rate of spherical particles under hindered settling conditions can also be calculated from Eqs. I and 2 by replacing p', the density of the fluid, by p" the apparent density of the suspen¬sion. The concentration of particles in the fluid thus imparts an appar¬ent density to the composite fluid or suspension greater than that of the liquid alone, resulting in a buoyant effect on the larger particles. Particle shape affects the settling rate of both coarse and fine particles. The general effect is to reduce their settling velocities and the effect is greater for coarse particles and for those settling under hindered settling conditions than for fine particles or free settling ones. For two particles of differing densities but settling at the same velocity under Newtonian conditions, the ratio of their diameters from Eq. 1, called the free settling ratio is: where L signifies the lighter particle and H, the heavier particle. Under Stokesian conditions the exponent would be 0.5. For hindered settling conditions the fluid density p' is replaced by the apparent density of the suspension, p", to obtain a generalized equation for the hindered settling ratio: assuming both particles settle in approximately the same regime. The free settling ratio as given by Eq. 3 has been called by Taggart88 the "concentration criterion" and is used to predict the effectiveness of any gravity concentration process (see Introduction to this section). Based on Eqs. 3 and 4, if two particles of densities pH and p,, settle at the same velocity, the diameter of the lighter particle will be larger than that of the heavier particle. For example, in the case of galena (pH = 7.5) and quartz (pL = 2.65) settling in water (p =1.0) under free settling, Newtonian conditions 3.9. Thus, a quartz particle nearly four times as large as a galena particle will settle at the same velocity. Any quartz particle just slightly less than four times the diameter of the largest galena particle may be separated from it. Under hindered settling, Newtonian condi¬tions in a suspension where p" = 1.65, dL/dH = 5.85 or any quartz particle just slightly less than about six times the largest galena particle may be separated from it. Reference to Eq. 4 indicates that a superior separation between two minerals of differing densities is favored by: (1) coarse particles settling under Newtonian conditions, (2) a large difference in (pa - PL), and (3) separation under hindered settling conditions where p" is high. Of course, there are practical limits to increasing p" excessively because at very high percent solids suspen¬sion fluidity would be lost and the hindered settling separation process defeated. Examples of hindered settling separators are the Dorrco-Fahren¬wald sizer,89 the Rheolaveur box 90 the Spitzkasten,89 and the Willoughby washer.91 These devices make a mineral separation on the basis of both specific gravity and size and all of them are essentially obsolete except for the Dorrco-Fahrenwald sizer and similar devices which still find application for the removal of coarse particles from a much finer particle assemblage and for preconcentration ahead of shaking tables. However, nearly all gravity concentration processes (jigs, tables, flowing film concentrators, heavy media separators) and many sizing devices (sizing classifiers, clarifiers, thickeners, hydrosepa¬rators) make use of the hindered settling phenomenon during the separation of particles. JIGGING Introduction In jigging, a mixture of ore particles, supported on a perforated plate or screen in a layer or "bed" with a depth many times the thickness of the largest particle, is subjected to an alternating rising and falling (pulsating) flow of fluid with the objective of causing all

Jan 1, 1985

By J. L. Thompson

The combustion of pulverized coal is a century-old practice. Many people have investigated coal dust and the residual fly ash particles that re-main after combustion. But the combustion process and fly ash formation mechanism are not under-stood. Working at our Maysville, Kentucky lime plant, the Dravo Lime Company Research team has collected samples of partially burned coal dust from the flame of a pulverized coal burner. Specimens were placed in a scanning electron microscope (SEM) equipped with an energy dispersive spectrometer (EDS). Microscopic examination of the particles, together with a large collection of micrographs of lime dust fused to fly ash particles, now make it possible to theorize as to how combustion advances through a piece of coal dust. Coal is a physical mixture of carbon-based compounds interspersed with bits of ash-forming materials, primarily alumina-silicate compounds. The alumina-silicate materials are commonly in angular nodules measuring 100 A to 700 A across their major dimension. It would appear that if the coal particle is in a well mixed gas stream, combustion starts at the surface and advances radially toward the particle center. With poor gas mixing, it is common that combustion moves from one end of a particle and advances on a broad front almost normal to some axis. In either case the combustion process and ash-forming mechanism are the same. As combustion occurs the carbon-based compounds oxidize in a highly exothermic reaction. Heat from the re-action melts the alumina-silicate nodules causing them to flow together and coalesce into an open sponge-like structure. As melting continues surface tension of the liquid and free surface energy causes the sponge-like structure to simplify into fewer, larger pores. Most likely the lighter molecular weight hydro-carbons burn before the heavier species. This means that as the melting ash structure is metamorphosing to a single large pore, there continues to be out-gasing throughout the structure due to combustion. Surface tension will form a spherical piece of fly ash if the ash material is liquid for a sufficient period of time. Out-gasing from combustion of the heavier hydro-carbons will provide internal gas pressure to keep the particle spherical until it freezes.

Jan 1, 1978

By Alexander N. Winchell

The cities of St. Paul and Minneapolis are located side by side on a double loop in the Mississippi River near the eastern boundary of Minnesota. Both cities lie on both sides 'of the Mississippi River, which in this region describes an S-shaped curve. The two cities together have a population of over 600,000 people, gathered in an area which has grown to be a single urban district inasmuch as the two cities have expanded until they occupy all the territory formerly between them, though the business centers are ten miles apart. St. Paul is located at the head of navigation on the Mississippi River and Minneapolis, farther upstream, at the Falls of St. Anthony. A recent water¬power development below the Falls of St. Anthony, constructed by the Federal Government, has at the same time made possible navigation to the foot of St. Anthony Falls through a lock around the dam which replaces the former rapids. Minneapolis, in 1920, had a population of 380,000 people, and St. Paul, 235,000. The two cities occupy land at an elevation of 700 to 900 feet above sea level. The average elevation is slightly less at St. Paul. Minneapolis has nearly 4,000 acres of parks and park¬ways, while St. Paul has nearly 2,000.

Jan 1, 1920

By Luis Yacher

CONTAC Engineers is a Chilean based company which works on process control and automation, principally for the minning industry. Jointly with other local and international engineering firms, CONTAC has participated in most of the last expansion and modernization plans being carried out at the Chilean pinning industry. At presente, CONTAC is involved on a research and development plan for the design and implementation of new instrument and process control procedures for mineral process automation. One of these efforts is the design of an ore size analyzer for coarse ore being transported over a conveyor belt. A feasibility and prototype study was made at a SAG Mill Plant and currently a high resolution, high speed analyzer is being developed. It will became available on early 1986.

Jan 1, 1986

By J. Andersen

The very large quantities of energy used for size reduction in the minerals processing industry necessitate the accurate prediction of comminution machine power draw for optimum plant design. The JKMRC has developed effective models for predicting ball mill, autogenous mill, semi- autogenous mill and crusher power draw. When combined with comminution models they provide a powerful tool for evaluating not only the size reduction performance of a plant design, but its power utilization efficiency as well. This paper describes the structure of the mill and crusher power models. Data from a wide range of equipment types are presented which illustrate the accuracy of these models.

Jan 1, 1992

By Danny Chan, Shahram Tafazoli, Nicholas Himmelman, Edmond Chow

"While wearable technology has been garnering attention in the consumer market, enterprise and industry have just begun exploring its potential applications. Since February 2014, Motion Metrics International has been working in collaboration with Vandrico Solutions Inc. to apply wearable technology to improve productivity and safety in the mining industry.This article discusses the value proposition of using wearable technology to deliver critical notifications in mining. The advantages and disadvantages of various form factors are explored. The challenges, both specific to mining and in general to wearable technology, are defined. Finally, the concept of MetricsGear, the first wearable technology platform developed for mining by Motion Metrics International in collaboration with Vandrico Solutions Inc., is described. MetricsGear is designed to notify the user with minimal distraction to keep the user focused on his or her primary task and aware of mission-critical events. This focus and awareness allows the user to work more efficiently and safely. The mining process consists of a series of intricate stages, where certain events or conditions in one stage can have a huge impact on other stages. The notification of these critical events or conditions is vital for ensuring the overall efficiency of the mining process. Critical notifications can be classified as two types: timecritical events and contextual information. Time-critical events are situations where a timely response from a responder can have a significant impact on seizing an opportunity or mitigating cost. Some human-computer interaction (HCI) experts classify notification types based on the interruption, reaction and comprehension (IRC) framework. Time-critical events are classified to have high interruption, high reaction and low comprehension due to their urgency. The responder should have just enough information to identify the crisis and react as soon as possible (Chow, 2014). For example, when a tooth is lost during shovel operation, it is necessary to recover the missing tooth before it jams the primary crusher. The consequence of missing this critical time period is a jammed crusher that can cause the mine to be out of production until the jammed tooth is extracted (Baumann, 2013). In this case, the responder only needs to see an image to confirm if the tooth is missing and information to identify which equipment and when the tooth went missing (Fig. 1)."

Jan 1, 2014

By D. L. Romero

The Line Creek Mine is operated by Crows Nest Resources Limited, a wholly owned subsidiary of Shell Canada Resources Limited. It is located approximately 25 km north of Sparwood in southeast British Columbia (refer to figure 1). The Line Creek coal deposit is part of the Upper Elk Coalfield in the East Kootenays. The coal is a medium volatile, low sulfur, bituminous coal and is used as a coking coal. Coal close to the outcrop is typically of low F.S.I. and is processed for thermal use. The present pits have a total coal reserve of 30 million clean tonnes and is a multi-bench, open-pit operation. The coal is recovered by front shovel and backhoe combination, and hauled to the plant with a fleet of 120-tonne and 40-tonne capacity trucks. Line Creek is unique in that it can simultaneously treat both metallurgical and thermal coal in separate processing facilities (refer to figure 2). Run-of-mine coal is delivered to the truck dump. The plus 50 mm material is scalped off and reduced to minus 50 mm in two rotary breakers. The raw coal is then conveyed to one of five, 5000-tonne capacity, silos. Typically one silo is used for thermal coal and four for metallurgical coal (refer to figure 3).

Jan 1, 1991

By Stephen P. Winkelmann

The expansion of gold production in Nevada's Carlin Trend over the last few years has been staggering. To increase gold production while minimizing capital requirements and providing for the safe and efficient mining and processing of ores has required a tremendous effort by the mine engineering group at Newmont Gold. Organizational changes were implemented by the company to effectively deal with a full range of engineering problems. Specially developed computer software and its applications will be explained in this article as well as improvements in the techniques used in converting geologic resources into reserves. Scheduling procedures for the mining of multiple pits to feed multiple processing facilities will likewise be presented. Ore control procedures and current mine engineering practices are discussed. Growth at Newmont Gold Expansion of production levels at Newmont Gold Co. was brought about primarily by the economic incentive to exploit increased reserves. From 1987 to 1989, mined production increased by 260%, gold production by 238%, and manpower by 166% (Table 1). These increases came quickly and demanded

Jan 1, 1990

Relatively few longwall mines in the U.S. operate under multiple seam conditions where the two seams are less than 200 ft apart. This paper describes the experience of six that do. These operations are located in PA, WV and UT, and include examples of both undermining and overmining. Some operate above or beneath their own workings, others are in historic mining districts and must contend with abandoned mines that are decades old. The lessons that these mines have learned cover a broad range of topics, including: • Whether to stack gateroads or place them under old gob areas; • How to size pillars and select artificial support to cross longwall stop lines; • How to use yield pillars to minimize multiple seam stresses and coal bump potential, and; • When to anticipate the creation of pathways for gas, water, or oxygen between current and abandoned gobs.

Jan 1, 2007

By T. P. Weldum, W. Conger

"The objective of the study conducted by the authors is to determine what media size(s) addition will maximize any given plant ball mill’s grinding efficiency (Fig. 1). The functional performance parameters “mill grinding rate through the size of interest,” and “cumulative mill grinding rates” from both plant and smallscale tests are applied to this task. A plant media sizing methodology, and industrial case studies, are provided.BackgroundA previous article (“Ball mill classification system optimization through functional performance modeling,” Nov. 17, McIvor et al., 2017, Mining Engineering) described circuit classification system efficiency (CSE), equal to the percentage of coarse (plus circuit P80 target size) material in the ball mill. It can be measured and then increased through pump and cyclone adjustments. By so transferring the application of ball mill power from fine (minus P80 product size) to coarse material, circuit efficiency is increased and overgrinding is decreased.The other factor that determines overall circuit efficiency is how well the power that is applied to the coarse material grinds it. This is affected by design and operating variables like the media sizing and mill percent solids. From circuit functional performance (McIvor et al., 1990):Circuit production rate of new fine material (t/h) = Power applied to coarse material (kW) x Mill grinding rate of coarse material (t/kWh) (1)The power applied to coarse material is the total mill power multiplied by the CSE. Therefore:Circuit production rate of new fine material (t/h) = Total mill power (kW) x CSE (%) x Mill grinding rate of coarse material (t/kWh) (2) Production rate, mill power and CSE can be measured during a plant circuit survey. The mill grinding rate of coarse material is then calculated. At the circuit target P80 size, this is the mill grinding rate (of coarse material) through the size of interest. The objective is to maximize this rate. The mill grinding efficiency can also be calculated by taking the ratio of the mill grinding rate over the material’s grindability as measured in the lab. Mill percent solids optimization has thusly been achieved (McIvor et al., 2000). For media sizing, maximizing the mill grinding rate on a given ore is equivalent to maximizing the mill grinding efficiency. Ores of different grindability will be part of the discussion."

Jan 11, 2018

By Dan B. Kivari

One of the key operations in any milling circuit is grinding. Over the last two decades, large diameter Semi-Autogenous Grinding (SAG) mills have become the preferred primary grinding mills. The drive systems became more complex as the size of the mills became larger, during which time operating and design experience for SAG mill accumulated. Considerable reliance is now placed in power electronics to provide flexibility in SAG mill operation to meet the demands of lower operating costs. Whereas earlier most SAG mills were designed to operate at fixed speeds, present day SAG mills can operate at a fixed speed or variable speed through the use of an AC synchronous motor or DC motor. However, this flexibility comes with an increase in capital and operating costs. Since the capital and operating costs of a project directly affect the project's rate of return, the selection of a SAG mill and its drive system should depend, in part, on the impact that the extra cost makes on the project. This impact can be measured by performing an economic analysis based on Net Present Cost (NPC) calculations. Although the NPC calculation is an important factor in the grinding mill selection, other factors such as reliability of mill drives and power source are important elements in the decision making process. This paper examines how the NPC calculation, plus reliability and power source, influence SAG mill selection.

Jan 1, 1991

By J. D. Miller

Fluid flow phenomena and therefore the flotation efficiency of air-sparged hydrocyclone (ASH) flotation are strongly dependent on the operating and design variables such as air flow rate, slurry flow rate, underflow and overflow opening areas, percent solids in the feed, reagent levels, particle size, ASH length, etc. Time-averaged multiphase flow characteristics, as deter-mined by x-ray CT and described in a previous publication, are used to understand and explain the steady-state ASH flotation of quartz. A detailed parametric study of quartz flotation by a nominal 2-inch diameter air-sparged hydrocyclone (ASH-2C) revealed the influence of these variables on the flotation response which is discussed in terms of the multiphase flow characteristics. Thus, a phenomenological description of quartz flotation has been established to assist in the further development and utilization of the ASH technology.

Jan 1, 1995

By Ian S. Lowndes, Adrian T. J. Moll

With the current trends towards the concentration of extraction to fewer but higher production units in deeper seams, often of higher coal rank, the role of methane drainage ranges has become critical. As mines exploit seams at greater depth and more remote locations there has been a need to extend the lengths of methane drainage ranges and to increase the capacity of the surface pumps. This trend has highlighted the need to effectively simulate the operation of a methane drainage range in order to produce the optimum design and operation of a safe and economic system. This paper describes the development of such an analytical tool. The application of this computer model is illustrated by the analysis of the operation of the drainage system of an extensive mine. The paper concludes by considering the development of a [Knowlege Based or Decision Support System] which will assist the ventilation engineer in assessing the current performance of an existing drainage range and will suggest possible remedial modifications or design strategies which may improve the system performance. INTRODUCTION In order to maintain the general body methane levels to below predetermined statutory threshold limit values (TLV's) on major longwall production units there are, for a given production rate, two major complementary control devices which may be employed by the ventilation engineer. These are the combined design and application of a competent mine ventilation and methane drainage system. The methane drainage system should be designed so as to maximise the capture of the methane (firedamp) from the seam being mined and the strata gas liberated in the vicinity of the extraction zone from gas bearing seams or other reservoirs of gas situated at above or below the mining horizon. The design fresh air quantity entering the face should ensure that under normal operating conditions the methane entering in the vicinity of the face is rapidly dispersed and dilute to well below statutory TLV's. The drainage system must therefore be designed to maximise the capture of the methane gas liberated to minimise the amount of gas which enters the ventilation airstream. GAS DRAINAGE The inherent dangers of methane emissions in mine workings have been a problem of concern to mining engineers for many years. To alleviate the risk, methane drainage has been employed for some years to capture the gas before it becomes a hazard to the operation of the mine. The method of drainage employed will be dependent on a number of factors which include the longwall mining method and production cycle employed, the characteristics of the coal seam being extracted and those of the adjacent gas bearing strata. The location of the gas bearing potential of the over and underlying coal measures must be determined. The gas bearing

Jan 1, 1993

By J. Pang, S. Chander

The mechanisms of chalcopyrite oxidation in alkaline solutions have been studied using cyclic voltammetry and impedance spectroscopy techniques. The cyclic voltammetry was used to elucidate oxidation/reduction reactions, whereas impedance spectroscopy was used for in-situ characterization of the interface. To analyze the impedance spectra, equivalent circuit models were constructed based on a layer- wise growth of reaction products. Parameters of the equivalent circuit were calculated by comparing the experimental spectrum with that expected on the basis of equivalent circuit. From the value of equivalent circuit parameters, properties of the surface layers were inferred. These properties change substantially with the change in oxygen and/or collector concentration in solution. From such changes the effect of oxygen and collector concentration on the contact angle and flotation behavior of chalcopyrite were interpreted.

Jan 1, 1991

By Joseph M. Zigich

The Anaconda Minerals Company operates a molybdenite and copper producing mine and mill facility near Tonopah, Nevada. Included with the milling facility is a ferric chloride leach plant for removal of copper mineralization from the molybdenite concentrates. Discussed are the engineering and construction schedules associated with the leach plant. The chemistry associated with the process along with a description of the process flow, equipment, and operating parameters are contained in the paper.

Jan 1, 1983

By Yoshio Hiramatsu

In Toyoha mine near Sapporo, the ground temperature is very high, perhaps due to inflow of hot mineral water, the origin of which may be rock masses intruded by the igneous activity in the Tertiary period. The rock masses are found about 2 km east of the mine, their temperature being estimated at from 770 to 870 K at a depth of 3 km. In this mine several veins bearing zinc and lead ores have been mined down to the depth of 300 m from the level of the main entry, of which altitude is 551 m above the sea level. Formerly no attempt was made at development of the deeper part of the mine than the 350 m level from the main entry. Let us indicate hereafter the levels in this mine in such a way that the level 300 m deep from the main entry is denoted as 300 mL. Since 1974, however, the 450 mL of Tajima vein has been developed by deepening No. 5 Shaft to, and by driving drifts in the level. Much hot mineral water has been flowing into the drifts from rock. After twenty-eight months from the beginning of development, the total length of excavation amounted to about 1300 m. The success in development so far may be attributed to cooling rock and air with chilled water and to adopting heat-resisting explosives and detonators. This cooling costed very much and for further development in this level a more effective and economical cooling is desired if possible. To obtain fundamental data for designing an economical cooling, the authors investigated, as a preliminary study, the heat balance in the main drift of 450 mL, by measuring the temperatures of air, rock, hot mineral water, the rates of flow of air and water and so on, in August 1976. They are now carrying out continual measurements of the variations in rock temperature, air temperature and humidity with time. The present paper describes the details of the preliminary study.

Jan 1, 1980