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By Thomas D. Henderson

Este capítulo presenta- un "caso histórico" de los pasos a dar para estimar IDS castes de capital y operativos para un concentrador típico de cobre porfirico de ± 9;070 toneladas métricas de producción diaria. El concentrador que se describe es hipotético, pero está basado en la experiencia propia y es, en efecto, una composición de varios proyectos reales. Las cifras utilizadas son ilustrativas solamente y no representan los castes reales de ningún lavadero en particular. Estas cifras son, sin embargo, representativas de los castes tipicos de molienda en un punto en el tiempo y son to suficientemente precisas coma para iIustrar los procedimientos de estimación de castes descriIos en este capítulo, No es posible publicar cifras reales y corrientes de los castes del lavadero y, en cualquier caso, las cifras de Ios castes reales pierden rápidamente su significación en Ia actual situación económica con los rápidos incrementos de las diversas partidas que configuran los castes. El trabajo está dividido en tres partes: (1) orden de la magnitud estimada, (2) estimación def estudio de viabilidad, y (3) comparación entre los castes estimados y los castes rea¬les. Con vistas a este trabajo supondremas que una mina tiene una propiedad minera explorada mediante perforación con testigos que refleja un yacimiento, ampliamente definido, como de unas 45,3 millones de toneladas métricas con una ley del d,6°/ de sulfuro de cobre, Esta compañia desea obtener primero una estimación del orden de magnitud de los costes (canto de los costes operativos como del capital) para determinar si la pertenencia garantiza más perforación y desmuestre de testigos. Suponiendo que esta evaluación económica obtenida utiIizando Ia orden de Ia magnitud de Ios costes sea favorable, sera a continuación requerido un estudio de viabilidad más preciso de los castes de capital y operativos. Seguidamente of recemas una descripción de los diferentes pasos dadospara desarrollar ]as costes del orden de primera magnitud y a continuación para los castes del estudio de viabilidad.

Jan 1, 1982

By D. J. Lachel, Douglas E. Hansen, Bruce Kennedy

The very high cost of exploration and development for new mines requires that a carefully planned, complete, and technically competent geotechnical evaluation be applied to the ore body and the surrounding rock. The ground conditions must be predetermined sufficiently so that, during the development and exploitation stages, unexpected conditions will be of minor consequence. Ground conditions are a major factor affecting the cost of excavated items such as shafts and inclines. For example, shafts cost less to sink in tight dry ground than in highly pervious ground below the water table, with all other factors being equal. The proper mining methods also is dependent upon the ground conditions. Bad ground for room-and-pillar mining may be ideal for caving methods; whereas, ideal ground for room-and-pillar mining would be extremely difficult for a caving operation. To do a proper engineering design, the parameters of strength, deformability, and permeability must be determined, in addition to the loads that will be applied as a result of the mining activity. The required degree of accuracy for each parameter depends upon the individual case. For example, the strength of massive rock in which openings of small cross-sectional area are to be driven at shallow depths needs to be resolved only within several thousand kilopascals (pounds per square inch). Under such conditions, the rock strength is not likely to be of great economic importance unless the rock is very weak. Large openings at great depths require a much more accurate appraisal of rock strength and deformation characteristics. Ground water conditions are important at any depth. Today, a number of techniques exist for evaluating conditions before opening up the ground for development. These techniques and their applications are described in more detail later in this text. The evaluation of ore body ground conditions should start at an early stage, as soon as the deposit has been proven viable for a mine. Economic geology investigations mainly serve to define the value and limits of the ore deposit. To do this properly, the geology of the site is determined in at least as much detail as is necessary to define the "controls" for the ore body. The ground-condition evaluation must determine the engineering properties of the ore body and the surrounding material. This includes the structural geology as it relates to rock strength; i.e., the frequency, extent, and orientation of discontinuities such as faults, joints, bedding, etc. Ground water conditions also must be determined. It is important that the detailed geology of the site be determined at an early date. Reasonable geotechnical parameters must be established during this time period, so they can be used in mining model studies that include both physical and mathematical models. The ultimate goal bf-the geotechnical evaluation is to predict the ground behavior as the excavations are made and how the ground behavior will affect the safety and economic well-being of the mine. The formulation of a ground-condition evaluation program should be planned carefully to determine the particular geotechnical parameters only in as great a detail as will have a bearing on the mine design. For instance, it is not necessary to make numerous triaxial tests in strong rock when the mining will be conducted within a few meters (feet) of the surface in a regime of gravity loading and an absence of tectonic stresses. Similarly, it is not necessary to invest large amounts of time and effort in the determination of joint strengths if the orientations of the joints are such that the failure could not occur along them. Generally, the determination of ore body ground conditions can be divided into four main steps: 1) The general and structural geology of the site must be determined; this is determined primarily during the economic evaluation. 2) The material behavior characteristics of the deposit and the vicinity must be determined. 3) The ground water hydrology of the deposit and the vicinity must be determined. 4) Miscellaneous ground support considerations must be established. Gross features are defined first, followed by a series of increasingly detailed investigations. Specifically, the geology study should start by utilizing techniques such as remote sensing, topographic map interpretation, geologic mapping, etc. After the gross features, such as major faults, have been delineated, the studies are directed toward defining the more detailed elements of jointing, rock strength, ground water hydrology, and other characteristics. This information is used to formulate an assessment of the strength of the ore body and the surrounding rock. Finally, all of the gathered information is integrated to make an assessment of the rock behavior in relation to the types of openings, the excavation methods, the support requirements, etc., that should be used for the rock conditions that are present. The mine design and the geotechnical assessment should be an iterative process. General conclusions about the rock conditions will dictate a type of mine plan. The specific mine plan then should be modified by specific geotechnical considerations. For example, an initial assessment may conclude that the ore body is strong and that a strong rock unit is present over the ore body. Thus, one definite possibility is to mine the deposit by room-and-pillar methods. However, subsequent investigations may reveal faults through the deposit, requiring relocation of the mine openings so that pillars and other structures are in sound rock. In other words, the final mine plan should be compatible with the geotechnical aspects of the ore body.

Jan 1, 1982

By William Lyman

GENERAL DESCRIPTION Shrinkage or shrinkage stoping refers to any mining method in which broken ore is temporarily retained in the stope to provide a working platform and/or to offer temporary support to the stope walls during active mining. Since ore "swells" when broken, it is necessary to shrink the muck pile a corresponding amount by draw¬ing some of the broken ore out as the stope is advanced-hence the name. Broken ore retained during stoping is drawn out after the stope has reached its limits. The stope may be left empty or may be filled with waste contemporaneous with, or subsequent to, the final draw. Traditionally the method implies conventional overhand stoping methods with miners working between the muck pile and the stope back, in a space which advances updip with mining and is maintained by balanc¬ing "swell" with "shrink." The shrinkage classification is also applicable to so-called "semishrinkage" methods in open pillar-supported stopes where broken ore is temporarily retained as a working platform but offers no wall support; and to various blasthole shrinkage methods which utilize broken ore temporarily retained in the stope for wall support, but which do not require miners to work from muck pile in the stope. The method is generally applied to steeply dipping veins of strong ore between strong walls. APPLICATION Geometry The geometry of a shrinkable vein is described in terms of dip, width, and regularity along dip. Overall strike and dip dimensions and irregularities along the strike generally impose no restrictions on the method. Dip is ideally 1.2 to 1.5 rad (70 to 90°). As dip falls below 1.2 rad (70°), the shrinkage draw begins to strongly favor the hanging wall side, thus leaving a poor working platform for conventional overhand work. This is particularly true in relatively wide stopes. The sup¬port afforded to the hanging wall also diminishes with decreasing dip, reaching nil as the dip approaches the repose angle of broken ore. Dips below 0.78 to 0.87 rad (45 to 50°) are not generally shrinkable except by open stope "seinishrinkage" methods. Minimum mining width is fixed by working space requirements in the stope-generally about 1 m. Shrink¬age in narrower veins requires that waste rock from one or both walls be broken with the ore and the attendant dilution accepted to achieve the minimum width. Nar¬row stopes are less suitable, encouraging hang-ups and bridging of broken ore, with the attendant problems of erratic draw and incomplete recovery of broken ore. Maximum practical width may be 3 m or less to over 30 m, depending upon the competency of the ore and its ability to stand unsupported across the stope back. This is a vital safety consideration in conventional over¬hand stopes, but is much less of a factor in blasthole shrinkage methods. Very wide veins and massive ore bodies have been mined by transverse vertical shrinkage panels separated by transverse vertical pillars which are either abandoned or recovered later by other methods. Regularity along the dip is a prerequisite of shrink¬age as there must be no serious obstruction to the flow of broken ore downward through the stope to the sill level. Gentle rolls along the dip are acceptable if the local footwall dip everywhere exceeds 0.78 to 0.87 rad (45 to 50°). Off-dip hanging wall and/or footwall splits can generally be mined selectively from a conventional shrink stope as they are encountered without ad¬versely affecting subsequent continuation of shrinkage mining updip on the main vein. Vertical offsets or major rolls along the dip which cannot be "smoothed over" generally require that a sublevel be established with new draw control development. Blasthole shrinkage methods are much less flexible (and thus less selective) in their ability to accommodate any of these irregularities. Ground Conditions The wall rock must be strong enough to stand with the minimal support afforded by the dynamic mass of broken ore in the stope. During active mining, local sloughing from the walls is restrained, but the broken ore affords little, if any, useful resistance to closure of the stope walls. Such squeezing, if present, may bind up the stope and cause the loss of much ore. Pillars left between and/or within stopes are effective in preventing closure but reduce overall recovery. Walls may be re¬inforced by bolting after each stope cut in conventional shrinkage but not in blasthole shrinkage. Ore in place must be strong enough to stand with no natural support across the stope width, although tem¬porary artificial support or reinforcement may be used locally in conventional stopes. Some spalling or sloughing is permissible in blasthole shrinkage as men are never present in the stope. Physical and/or mineralogical characteristics of the broken ore may impose restrictions on stope design and/or operational plan¬ning, and may even preclude the use of shrinkage al¬together. Examples include: ores which, when broken, are cohesive or which tend to pack or cement together under the influence of ground water, wall pressure, and/ or chemical reaction. Such conditions precipitate er¬ratic draw during mining and often result in difficult and/or incomplete final draw; pyritic ores which oxidize very rapidly in the stopes and may generate heat, imposing a fire hazard by spontaneous combustion; sulfide ores which oxidize sufficiently in the stopes to adversely affect mill recovery by flotation; and ores (es¬pecially those containing uranium minerals) which ex¬ude radon gas and thereby impose ventilation constraints on stope design. In most cases these problems can be minimized by limiting the size of stopes, by minimizing the duration of mining activity in each stope, and by promptly drawing each stope empty following comple¬tion of mining.

Jan 1, 1982

By J. P. Van Dillen

Continuing interest in processes that will effect separations in the fine particle size range has prompted intense interest in the use of high gradient magnetic separation (HGMS) for beneficiation of uranium ores. Some of the uranium minerals are paramagnetic and hence are amenable to concentration by HGMS. This paper presents a brief review of the problems associated with uranium ores beneficiation, a description of the HGNS process, and the effects of the parameters--pulp velocity, field strength, matrix loading and magnetic field gradient on product recovery and grade. Unsatisfactory separation of uranium minerals by HGHS is related to ore specific properties such as lack of liberation, fine dissemination of uranium minerals, uniform distribution of magnetic minerals throughout the ore matrix and interference of micron sized slimes. Experimental techniques to identify and resolve some of these problems are outlined.

Jan 1, 1982

By Richard Klimpel

The laboratory ball mill grinding of coal/water slurries is described for three coals. Emphasis is placed on the measurement of specific rates of breakage for various operating conditions. Corresponding rheological information is also presented based on data collected using a rotational viscometer. A rheological explanation of the influence of various slurry fluidity conditions on the specific rates of breakage is presented. This rheological approach has proved useful In both the identification of suitable chemicals for increasing breakage rates in coal/mineral slurry grinding and in the plant scale implementation of dense slurry grinding systems.

Jan 1, 1982

By Jalal Vakili

In designing very large excavated slopes which are becoming increasingly common in both mining and civil engineering projects, the most important decision of the engineer is the selection of a slope angle. The purpose of the geotechnical study of the slope is to ensure reasonable stability of the slope in the most economical way. Because the rock mass composing each slope is unique, there are no standard solutions for slope stability analysis. A practical solution is formulated from the basic geologic data, engineering properties of the rock, geometry of the slope, and ground water observations. Stability analysis consists of analyzing the forces causing and resisting slope failure. There are several methods of slope stability analysis. The limit equilibrium method evaluates the overall stability of the sliding mass just on the verge of slip, using some or all of the three equations of static equilibrium of a plane problem. The stress-strain relationships of the rock mass are not considered. Method of analysis include the assumption of the shape of the slip surface. Peterson (10) introduced the Swedish method of analysis in 1916, which is based on cylindrical sliding surfaces. Fellenius (4) in 1926 extended this method and introduced the method of slices. This method was refined by Bishop (I) in 1955 and later used by Morgenstern and Price (8) in 1965. In many cases layers of extremely soft strata or strata weakened by neutral stress cause failure to take place by translation along plane of weakness instead of by rotation on a circular surface. In this paper, a slope with a horizontal layer of weak meterial is selected for analysis. This is a usual case for open pit lignite mines where lignite seam may be non-cohesive and 'weaker than the overburden rock or a soft clay layer underlies the seam. It is also a common case of rock fill dams supported on a foundation stratum of clay where the strength of the clay layer is less than that of the rock fill. It is shown that utilizing this method, the location of potential failure surface and the value of the minimum factor of safety can be calculated directly and the trial computations of factors of safety for a number of assumed failure surfaces are not necessary.

Jan 1, 1982

By Thys B. Johnson

This paper presents a new algorithm for solving the Controlled Flow Mine Ventilation Problem in an optimal manner - minimum power consumption. The use of booster fans and restrictions on the locations of ventilation regulators are dealt with in a straightforward manner. This new approach is based on a common Operations Research algorithm for which there are many fast computer programs available.

Jan 1, 1982

By Richard L. Bullock

At the beginning of the previous chapter, variation in the types of room-and-pillar stopes were briefly identified so that the reader could begin to understand the extensive application of this mining system. These same variations will be reviewed and elaborated upon in this chapter while discussing the methods by which the material is extracted. FULL-FACE SLICING BY DRILLING AND BLASTING When the entire mineralized thickness is taken in one pass of mining, it is known as full-face slicing. There is no mineral of economic value intentionally left either in the floor or the back. Theoretically, there is no limit as to how high the face could be in a single pass. But there are practical limitations to equipment size, pillar stability, and control of local loose slabs on the face, roof, and pillars that dictate current practices. The range of face height of the 15 typical room- and-pillar mines covered in Table 8 (Sec. 1, Subsec. 7) from the Dravo report, 1974 was 1.7 m (5.5 ft) to 9.8 m (32.1 ft), the average being 5.2 m (17.1 ft). In the Gaspe copper mine, Murduchville, Que., the drill jumbos were constructed with an extendable tower for full-face mining up to 15.2 m (49.9 ft) if necessary. However, a more common practice at that mine is to take the ore in multiple slices (Hall, 1959). In many metal mines the normal practice when starting stoping is to drive a single development drift into the ore zone a distance that will allow at least four or five rooms to be opened on each side of the initial drift. Since this opening will probably serve as the main haulage drift, it should be kept as straight and as level as possible. As to the other criteria for the stope developments, it might be well at this point for the mine planner to review the remarks presented earlier (Sec. 1, Subsec. 7) on drifts, entries, and crosscuts for production. After the initial drift is driven, it should be slabbed to the full room width if it is not that wide already. Next, the future pillars should be marked on the ribs and the rooms driven between them. If the ore extends beyond the length of the initial drift and is to be mined in the normal sequence of mining, care should be taken in a random room-and-pillar system to see that the extension of the initial roadway remains straight. If pillar "spotting" is left to the shift foreman or to the miners, they probably will not realize the importance of maintaining a uniform pillar line next to the roadway, and of keeping the line straight. Invariably, a pillar will end up in the middle of the future drift extension resulting in a "dogleg" in a possible main haul- age road. While this error is not catastrophic, too much weaving in and out of pillars will certainly slow down future production haulage and needlessly increase costs. Face or Breast Drilling and Blasting Practices To drill and blast the initial advances into the rock, usually some form of cut pattern must be used. In room-and-pillar mining, when a pattern is drilled in a face and that face is the only exposed surface to which all of the rock must be broken, the pattern is known as a "face round" or "swing." The face to which the rock must be broken is referred to as a "free face." Thus the most common way to advance a room into virgin rock with only one free face is by drilling swings. If, after breaking around a pillar, taking down back, or taking up bottom, there is a second (or third) face exposed, a group of holes drilled nearly parallel to a free face and breaking to it is known as a "slab round" or "slabbing." In some mining districts it is also known as "slashing." Obviously, because two faces exposed offer less resistance to fragmentation than a single face, slabbing requires less drilling and fewer explosives than required to break the same tonnage with the same degree of fragmentation with swings. Therefore, it behooves the driller (or supervisor) to plan the combination of

Jan 1, 1982

By Davies L, Jeffreys J

The Broken Hill Proprietary Company Limited (BHP) has been successfully using mini computers in its open cut strip and pit coal planning for the past six years. A recent development within the computerised Seam deposit User oriented Mine Planning System (SUMPS) provides the facility for the generation, evaluation, and simulation of colliery workings.Existing computer programs for collieries were studied, but none met the requirements of BHP's underground mine planners.Consequently, it was decided to develop programs, based on those existing within BHP for open cut strip and open cut pit mine planning. The new planning facility was designed to enable any number of proposed colliery developments and operating plans to be evaluated against optimistic, most likely and pessimistic geological lease area interpretations.In mine planning, sensitivity studies should be conducted to determine the likely effects on production, development costs and colliery economics of changes in mine design and budgeted equipment productivity. By automating the manual calculations involved, the computer programs provide the ability to conduct numerous and extensive sensitivity studies. Hence an optimum mine design for budgeting purposes can be derived quickly with an improved degree of confidence in the long term viability of the proposed development.

Jan 1, 1982

Over the past few years, the School of Mines of Paris has carried out a program of research* on the grade-tonnage relationships in the border zones of porphyry copper deposits. The initial problem was quite difficult : how does the grade vary at the border of the known orebodies? Does it decrease regularly, or do some thresholds exist, where the grade diminishes abruptly? The answer to these questions have important repercussions on the prediction of the future reserves in the fringe of the orebodies. If the grade decreases regularly, very important reserves could exist; if not, there would be strong limitations to the possible recoverable reserves. The first step of our study was a critical study of previous work done on the subject. It appeared that, following Lasky's pioneering paper (Lasky, 1950), most of the authors used what we can call an historical approach to the problem. In this, the grade-tonnage relationships are calculated from the available figures for production and reserves and the results are then extrapolated to the future. The strong limitation of this model are brought out in this article.

Jan 1, 1982

By J. Wong

We report a successful use of x-ray absorption spectroscopy to probe the chemical and structural environment of trace elements in coal. When applied to vanadium in coal, it was found that vanadium exists in at least two environments: a trivalent site 6-fold coordinated by oxygens at ca. 2.0~ and a tetravalent site 5-or 6-fold coordinated also by oxygens, one of which forms a short V=O bond at ca. 1.6a. No evidence of nitrogen (porphyrin)1'or sulfide environment was found. It was also found that the vanadium environ-ments in the raw coal did not survive unchanged in a liquefaction process. These findings have implications for coal cleaning processes and for trace element release into liquefaction process streams.

Jan 1, 1982

By K. P. Sinha, J. F. Schatz, T. F. Borschel, S. Demou

Triaxial creep behavior of oil shale was investigated in the laboratory at simulated in-situ conditions. A range of temperature and stress conditions were chosen to represent those within the inter-chamber pillars of a modified in-situ retorting operation. Oil shale in grades up to 60 gallons per ton have been tested. Experimental data and a constitutive model have been presented here.

Jan 1, 1982

Mining operations have been carried out at Mt. Lyell for the past century. Until the 1970's the location of the mines in relation to the surface topography has enabled the mines to drain their copious inflows of water by gravity. Deeper mining has forced the Company to undertake pumping of increasing inflows of water. A major pumping facility is being established near the base of the known ore reserves with a view to serving the future requirements of the mines. The system has a design capacity of 600 litres per second. Run of mine water will be treated in Vertical Flow Settlers to produce clear water and thickened mud, each of which will be disposed of separately. The clear water will be pumped by six 1000 kW conven- tional centrifugal pumps. The thickened mud will be removed by a Hydraulic Displacement System. The rising mains for both the mud and the clear water are stainless steel cased bore- holes.

Jan 1, 1982

By Christian F. Baiz

Industrial minerals can have similar financing requirements to those needs of metaliferrous and coal mining projects. Although, one may see many similarities in the fundamentals of the various risks that must be evaluated in lending to any specific project, one of the outstanding differences is the understanding of the many different markets that these industrial minerals represent. There are in excess of 30 principal industrial minerals ranging in terms of annual production from the largest quantity such as sand and gravel, to possibly the smallest quantity of production of a legitimate industrial mineral which may be defined as industrial and rough gem diamond. If we assume that it is necessary to finance independent projects of ore bodies of industrial minerals for development, the financing techniques are similar to the fundamentals that we would use for financing coal projects or metaliferrous deposits. Therefore, let's begin by redefining project financing in the traditional banking or private sector area, and then in the later development of this paper, we will apply these techniques to industrial minerals in general and cite it with several specific examples.

Jan 1, 1982

By Ernest W. Loya

A method for preparing gravimetric primary standards of helium in the parts per trillion range has been developed by the Bureau of Mines. The standards are used to calibrate a specially modified mass spectrometer used in determining the abundance of helium-3 and helium-4 in geochemical samples and of helium-3 in isotopically purified helium-4. The desired gas standards are prepared in 5-kg metal cylinders filled to 1,800 psia. Relative errors of less than 1 pct have been attained in the parts per trillion range using a successive dilution technique.

Jan 1, 1982

Topography, location of coal bearing land and existing built-up areas severely restrict locations for overburden disposal and dumping sites north of the Latrobe River clear of the coalfields are assumed until internal dumping areas can be established.

Jan 1, 1982

US 4,201,749-In the Bayer process production of alumina wherein aluminum hydroxide is precipitated from a sodium aluminate solution for use as seed, sodium oxalate and caustic components are removed from the precipitate by washing it first with a washing liquor which is saturated with sodium oxalate but has little content of caustic, and secondly with hot water. Substantially no sodium oxalate is recycled, yet loss of sodium aluminate is minimal. K. Yamada, T Harato, and H Kato, assigned to Sumitomo Aluminum Smelting Co Ltd May 6, 1980 8 pp. (423-127). Filed. 5-8-79 (37,145) Priority. Japan, 5-8-78 (53/54,803). Same¬British 2,020,642, dated Nov. 21, 1979.

Jan 1, 1982

By Mann LD

The Haveluck gold deposit, located 2.5km north of Meekatharra, contained 1,284,000 tonnes at an average grade of 1.8 g/t gold of drill delineated proved and probable ore prior to commencement of production. Following successful laboratory and pilot scale testing a decision was made to establish a 300,000 tonnes per year heap leaching operation. The ore, crushed to minus 100mm was stacked on an impermeable membrane and irrigated with a dilute cyanide solution. Gold was recovered from the resulting pregnant solutions by adsorption onto activated carbon and gold bullion produced in a modified Zadra process desorption plant. From the commencement of mining it was obvious that a high percentage of fine material was being produced which would adversely affect the percolation of the cyanide solution-through the leach heaps. A screening plant was incorporated into the crushing circuit which provided an oversize product satisfactory for heap leaching and an undersize for carbon-in-pulp treatment. Plant design provided for the slurrying of minus 0.8mm ore before feeding to the pachuca tanks. Initial practice was to use four tanks for gold dissolution and two for carbon contact. Vibrating screens were used through- out for the carbon-ore separation and the loaded carbon was stripped in the existing desorption plant.

Jan 1, 1982

By Kirby D. Mellegard, Ralph A. Wagner, Paul E. Senseny

Six creep laws for salt, each fitted to the same laboratory data base, are used in a numerical model of the deformation of an opening in salt to determine the influence of creep law form on the predicted closures. Results show that predicted closures can be very sensitive to the functional form of the creep law chosen to model the laboratory data.

Jan 1, 1982

Provide roof to floor closure rate measurements during underground mining, especially retreat mining of coal pillars, thereby helping the miners predict an imminent roof fall. Approach An electronic device and rugged potentiometric extensometer were developed to provide a means fur measuring total closure and the rate of closure. How It Works An intrinsically safe, battery operated instrument is enclosed in a rugged, hand carried field case. The unit is constructed using low-power, integrated circuits to provide minimal battery drain. Two digital Light Emitting Diode (LED) readouts are used; one displays the rate, the other displays the total closure. The unit will measure one extensometer continuously or two extensometers alternately. Two potentiometric extensometers with a closure measurement range of 6 inches are attached to the readout unit. As the roof to floor closure changes, the potentiometric resistance

Jan 1, 1982