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By Herbert Clark Hoover, Lou Henry Hoover

PREVIOUSLY I have dealt with the methods of separating silver from copper. There now remains the portion which treats of solidified juices ; and whereas they might be considered as alien to things metallic, nevertheless, the reasons why they should not be separated from it I have explained in the second book. Solidified juices are either prepared from waters in which nature or art has infused them, or they are produced from the liquid juices themselves, or from stony minerals. Sagacious people, at first observing the waters of some lakes to be naturally full of juices which thickened on being dried up by the heat of the sun and thus became solidified juices, drew such waters into other places, or diverted them into low-lying places adjoining hills, so that the heat of the sun should likewise cause them to condense. Subsequently, because they observed that in this wise the solidified juices could be made only in summer, and then not in all countries, but only in hot and temperate regions in which it seldom rains in summer, they boiled them in vessels over a fire until they began to thicken. In this manner, at all times of the year, in all regions, even the coldest, solidified juices could be obtained from solutions of such juices, whether made by nature or by art. Afterward, when they saw juices drip from some roasted stones, they cooked these in pots in order to obtain solidified juices in this wise also. It is worth the trouble to learn the pro- portions and the methods by which these are made. I will therefore began with salt, which is made from water either salty by nature, or by the labour of man, or else from a solution of salt, or from lye, likewise salty. Water which is salty by nature, is condensed and converted into salt in salt-pits by the heat of the sun, or else by the heat of a fire in pans or pts or trenches. That which is made salty by art, is also condensed by fire and changed into salt. There should be as many salt-pits dug as the circumstance of the place permits, but there should not be more made than can be used, although we ought to make as much salt as we can sell. The depth of salt-pits should be moderate, and the bottom should be level, so that all the water is evaporated from the salt by the heat of the sun. The salt-pits should first be encrusted with salt, so that they may not suck up the water. The method of pouring or leading sea-water into salt-pits is very old, and is still in use in many places. The method is not less old, but less common, to pour well-water into salt-pits, as was done in Babylon, for which Pliny is the authority, and in Cappadocia, where they used not only well-water, but also spring-water. In all hot countries salt-water and lake-water are conducted, poured or carried into salt-pits, and, being dried by the heat of the sun, are converted into

Jan 1, 1950

By P. Radziszewski

"The comminution energy curve is a profoundly simple graphical technique that displays detailed operating efficiency information for industrial comminution processes. The database underlying the methodology has recently more than doubled, to now contain a substantial fraction of the international comminution output (1.7 billion tonnes of rock per year) across more than 8 major commodities. This increase has deepened the analytical capabilities and effectiveness for assessing comminution circuit performance in relation to comparable operating mines. This paper will explore how the energy curve methodology can be used to examine the ability of stirred mill technologies to increase efficiency.Specifically, this paper explores how the energy curves methodology can be used to present energy efficiency gains achieved at three mines employing a stirred mill technology under different duties. The first of these cases studies was conducted by Rosa (2014) who compared the performance of two ball mills operating in parallel with a Vertimill in the regrinding circuit at the Samarco Iron project in South East Brazil. The Vertimill was found to achieve the same product specification while consuming 40% less energy. Brissette (2010) investigated the impact of grinding media on the performance of both ball mills and Vertimills. The energy savings achieved varied from 10% to 40% for the ball mills and 30% to 60% for the Vertimills. Lastly, Bergerman (2012) investigated the operating conditions of the Sossego regrind circuit from 2006 to 2011 against design specifications and laboratory testing. The operating work index of this circuit varied from 5.2 kWh/t to 44 kWh/t over this 5 year period. The full suite of energy curves"

Jan 1, 2016

By G. A. Swanquist, E. M. Morales

INTRODUCTION The idea of water management and control at the Church Rock Mill operations began to take shape in February 1979. At that time, we were already investigating the feasibility of decreasing the fresh water requirements so that the solids would become the limiting factor in tailings impoundment utilization. The area for solution evaporation could be kept at a fraction of the normal requirements under the standard process of full water usage. The Church Rock Mill is an acid leach circuit followed by solids/liquid separation with thickeners in counter current decantation, and solvent extraction. Following the normal design of acid leach circuits, reuse of tailings solution was not incorporated in the original mill process design. INITIAL WATER CONTROL INVESTIGATIONS The investigations to decrease the fresh water requirements centered around modifying the grinding circuit from the present semi-autogenous grinding (SAG) mill in closed circuit with hydrocyclones, to open circuit grinding with a rod mill. The open circuit grinding with the SAG mill and rod mill in series had the potential of decreasing the water requirements for grinding and leach dilution by approximately 50% or 1.4 m3/min (300 gpm). The grinding pulp density would be maintained at 70 to 72% solids, and the leach dilution to 50% solids would be accomplished with acid tailings liquor recycle. In such a grinding circuit arrangement, the SAG mill would provide the primary or coarse grind, and the rod mill would be used for the fine grind. By the SAG mill and rod mill series grinding method of water control and other secondary water controls in various places downstream from the grinding circuit, the required necessary evaporation area was estimated at 120 acres of liquid surface. A second method of water control at grinding was investigated. A two-stage cyclone classification circuit appeared to have a good potential of achieving the same water reduction at a much lower capital and operating cost. However, in retrospect, this would not have been a viable method since a high slime recycle load would have been established hindering classification. The use of reagents to neutralize the acid tailings solution was not considered seriously at that time, since it would have materially increased operating costs, although it would have also allowed more tailings solution recycle and consequently, less fresh water usage. However, with the tailings solution deposition area available at that time, it was not then necessary to incur the high cost of neutralization. The control expected by the series grinding of semiautogenous and rod mills would have been sufficient to maintain a water consumption/evaporation equilibrium well in line with the available land area. IMPLEMENTATION OF NEUTRALIZATION OPERATIONS During the summer of 1979, the UNC Church Rock Mill experienced a tailings dam breach which resulted in a prolonged mill shutdown. Upon resumption of operations at the end of October 1979, tailings deposition was restricted to a small portion of the tailings impoundment area. Figure 1 shows the general tailings area and the limits of the present deposition area in the central part including the borrow pits. These borrow pits had been excavated to provide materials for tailings dam construction. Immediately after resumption of operations, it became evident that it would be necessary to control the quantity of liquid to be evaporated because of the small confined area available for tailings solution deposition and to maximize the deposition time in the tailings area. The water control required had to be exercised on a large scale, and to be in operation as quickly as possible. An obvious solution was to reuse the tailings liquor in mill process. Immediate steps were taken to install the necessary equipment for tailings neutralization on an interim basis. Anhydrous ammonia was selected as the primary neutralization reagent since it was the quickest system that could be placed in operation. Previous laboratory tests indicated fair results with ammonia neutralization. Such a system required a minimum of installed equipment and handling. INITIAL NEUTRALIZATION OPERATIONS Actual neutralization operations began on November 26, 1979. The raffinate solution which normally would have been discarded was pumped to a 3.7 m (12ft) diam by 4.3 m (14ft) tank for reagent contact, see Figure 2. At this tank, anhydrous ammonia was added directly from the tanker trailers and controlled at pH 7.0 nominally. Agitation was provided by air sparging. The neutralized product formed a highly viscous slurry in the grinding circuit which resulted in pumping and cyclone classification problems.

Jan 1, 1982

By Richard Stoltz, Gary Smith

This paper presents the findings of a face ventilation investigation, conducted by the Ventilation Division, Pittsburgh Health Technology Center, at a underground coal mine operating in the Illinois No. 5 coal seam. The purpose of the investigation was to evaluate the effectiveness of a blowing face ventilation system in controlling methane liberation in the face area while mining with brattice distances up to a maximum of 40 feet from the face (extended cut). The continuous miner on this section was remotely controlled and equipped with a machine mounted dust collector. Data collected during this investigation included methane concentrations, intake and return air quantities in the face area, and scrubber air quantities. In addition, a time study was conducted to correlate the methane peaks to activities in the face area. In conjunction with the face study on Section A, separate methane liberation studies were conducted on three additional sections: B, C, and D. This was done in order to compare intake methane concentrations and methane liberations from other sections in the mine to the test section. Calculations indicated that the minimum intake air quantity necessary at the inby end of the line brattice for the section investigated was 7,000 cfm. Due to the higher intake methane concentrations and average methane face liberations measured on three other sections, the minimum air quantities necessary at the inby end of the line brattice were determined to be 13,900 cfm, 11,100 cfm, and 15,000 cfm, respectively.

Jan 1, 1991

By N LaBranche

The resurgence of Silicosis and Coal Worker’s Pneumoconiosis (CWP) in coalmines has placed the spotlight back on the management of worker exposure to particulate matter across all mining sectors (Coal Workers Pneumoconiosis Select Committee, 2017). There already exists a substantial body of knowledge on particulate matter and its impacts upon human health (National Academies of Sciences, Engineering, and Medicine, 2018). The contribution of silica to respiratory disease is well known and there is a growing concern for exposure to diesel particulate matter. However, it is also clear that there are major gaps in our understanding. In part, those gaps relate to the contribution to adverse health effects of mineralogical constituents in the dust (LaBranche et al, 2021a). It is also becoming evident that the size and shape of the particulate matter can affect its potential impact on human health.The University of Queensland has developed a methodology for characterising respirable and inhalable dust samples using scanning electron microscopy. Multiple samples from numerous mining operations have already been analysed using the UQ methodology. The process has been successful in showing the variety of mineralogica lcomponents and particle size distributions presentin various areas of the mines.

Oct 12, 2022

By Khandaker M. Anwar Hossain

This paper presents the development of lightweight concrete (LWC) incorporating pumice aggregate and volcanic ash (VA) based ASTM Type I blended cement (PVAC). Fresh and mechanical properties of LWC mixtures such as slump, air content, compressive strength, tensile strength, density and modulus of elasticity are described. The durability characteristics are investigated by drying shrinkage (DS) and water permeability. The investigation suggests the production of LWCs incorporating blended PVAC and VPA having satisfactory strength/durability characteristics for structural applications. The use of PVAC induces the beneficial effect of reducing drying shrinkage and water permeability. Development of such non- expensive and environmentally friendly LWCs with acceptable strength and durability characteristics is extremely helpful for the sustainable development and rehabilitation of volcanic disaster areas around the world.

Jan 1, 2010

By M. E. Austin

Many numerical models for the tubular solid oxide fuel cell (SOFC) exist, with various degrees of sophistication. However, the question of what physics must be modelled, and what can be neglected has not been resolved. A parametric model is proposed and analyzed to quantify the influence of thermal phenomena such as reaction heat generation, cathode channel convective cooling, and thermal radiation with participating media. Currently, a three-dimensional model is used to simulate the conditions of an anode supported cell in the centre of a tube bundle. The interconnect and nickel foam spacers that bridge adjacent cells in series and parallel, respectively, are neglected thus enabling a domain consisting of a 45o azimuthal section of the tube. The conservation equations for mass, energy, and species transport are solved using commercial computational fluid dynamics software, and radiation heat transfer is also modelled using a Discrete Ordinates method. Finally, temperature-dependent material properties and reaction rates are used. The model predicts lower overall temperatures than models that do not include radiative heat exchange and more moderate temperature gradients than those that ignore participating media radiation. Further, different regimes are identified, from convection dominated heat transfer to radiation dominated heat transfer. The results can be used to aid in the selection of appropriate models for a given fuel cell operating regime.

Jan 1, 2005

By Y. Topkaya, R. E. Johnson, B. R. Palmer, R. B. Bhappu, J. Brent Hiskey, Gürses Öner

Currently there is considerable interest in applying in situ leaching technology to recover copper from oxide ores. In many instances, the distribution of copper mineralization in these deposits is fracture controlled. Laboratory core-leaching experiments have demonstrated that leaching response is sensitive to the dissolution of copper from and the formation of channels along main fractures. Solution short- circuiting can be a major limitation of this technology. This paper describes the application of macromolecular polymers to achieve flow-profile control during leaching. Pre- and post-leached fractures were characterized by computerized tomography (CT) analysis.

Jan 1, 1993

By Aladar B. Dory

INTRODUCTION Practical experience in mines with known presence of radon daughters in the mine atmosphere in Canada and elsewhere shows that a very high concentration builds up in an unventilated dead end heading. As Holaday et al1 observed, even a minimal air movement results in a drastic reduction in radon daughter concentration. It is therefore obvious that the main objective of radon daughter control in the working environment is to design the ventilation system providing an optimized flow of fresh air into the workplace, resulting in acceptable climatic conditions and achieving radon daughter concentrations resulting in exposures as low as reasonably achievable. BASIC OBJECTIVES Large mining companies, having extensive material resources and professional expertise, have utilized elaborate electrical modelling in the design of mine ventilation systems as early as 1950 (coal mining industry in Europe) and with the advance of computer modelling techniques, their utilization in ventilation systems design is on the increase. Unfortunately, these methods are usually not available to small mining companies and even the large companies might not achieve the fullest benefit from utilizing them, if proper limiting factors are not considered in the modelling. When an evaluation of a ventilation system of a mine is undertaken in literature, a measure of the amount of air supplied underground per one ton of ore mined is used as an indicator of the efficiency of the ventilation system. Yet, even the greatest amount of air forced into the mine might not result in an acceptable working environment if a proper distribution of this air into individual working places is not achieved. The volume and the age of the air are probably the two most important factors in achieving acceptable radon daughter concentrations in the workplace, but other factors also have to be considered. DIRECTOR MINE - ALCAN, NEWFOUNDLAND FLUORSPAR WORKS ST. LAWRENCE, NEWFOUNDLAND, CANADA Ventilation To illustrate the effects of the design of the ventilation system on the control of radon daughter concentration, let us review the gradual development of the ventilation system of this mine from the earlier years of its development up until its final years of operation. This mine, located near the community of St. Lawrence on the south coast of Burin Peninsula was developed in the late thirties and reached full production by 1942. Unfortunately as was customary at that time, the only source of ventilation was a natural draft. The mine was extremely wet, and no significant attention was initially given to possible health effects of dust. It was not until the mid-fifties, when a number of cases of silicosis had surfaced, that de Villiers and Windish2 observed a significant increase of lung cancer incidence among the miners in comparison to its incidence among the general population of Newfoundland. Suspicions regarding radiation as a cause of the lung cancer were expressed, but it was only in surveys taken in late 1959 and early 1960 that Windish3 and Little4 established the presence of radon daughters in the mine atmosphere in very high concentrations. Windish, de Villiers and Hurley suggested that the most likely source of the radon in the mine was the mine water which dissolved radon during its passage through the granitic country rock in the surrounding geological area. This conclusion was confirmed by analyses of water from various areas of the mine by the Atomic Energy Canada Limited laboratories. The radon values in the samples varied from 4,240 to 12,850 pCi/L5. Following the discovery of the presence of radon daughters in the mine, the company took speedy action to install mechanical ventilation for the mine. The system was not designed as a total unit, but fans were installed rather on a trial and error basis. The basic system installation began in March 1960 and was completed by 1962. It remained basically unchanged with only minor modifications until August 1973 when a wholly new, redesigned ventilation system was implemented. A schematic section of the mine and its ventilation system for the period prior to March 1960 is given in Figure "A", for the period 1960-1973 in Figure "B", and for the period after August 1973 in Figure "C". The ventilation system prior to 1960 is not known. All workings of the mine were ventilated only by natural ventilation. If any measurements of airflows at different or any times of the year ever existed, no records have been preserved. The very minimal natural ventilation was augmented by "blowing" air from compressed air supply lines and exhaust air from drills. It is known that the compressor capacities of the mine were limited and therefore no significant air movement was probably created by the "blowing".

Jan 1, 1981

By Gordon R. Wicker, Peter V. Avotins, Steven S. Ahlschlager

Laterite ore slurries exhibit Bingham plastic behavior and can be characterized by the magnitude of their yield point and Bingham viscosity. These two parameters are affected by ore composition, solids concentration, and temperature and are used to predict the pressure drop in pipelines and heat transfer coefficients in shell-and-tube heat exchangers. In cycloning of laterite slurries, the calculated cut size is corrected by factors obtained from the rheological parameters.

Jan 1, 1979

By R. B. Shu

A moderate thermophilic microorganisms, has been isolated (45?) from a typical sulphide mine in the region of Mo Jiang located in the southwest of China. The culture has been used to extract copper from chalcopyrite concentrates obtained from De Xing copper mine. The majority of experiments were carried out in 300mL Erlenmeyer flasks containing 150mL volumes of mineral suspension. Effects of some variable parameters such as solids concentration, pH and the concentrations of inorganic nutrients on the bioleaching of chalcopyrite obtained from De Xing copper mine were investigated. For a given concentrate the most important factor affecting the fractional copper extraction was the size of the mineral particles. The overall rate of leaching was found to be limited by the area of mineral surface. The attack started in cracks and genetic defects and finally reached the whole chalcopyrite surface. As a consequence of this, the ore showed strongly cracks. The additive of ferrous sulfate increases the bacteria activity, but did not improve chalcopyrite bioleaching. The results indicate that the nature and relative proportions of the various minerals present have been shown to be important factors.

Jan 1, 2014

By M. J. Taffanel

I AM very much impressed by this manifestation of international brotherhood; the mining engineers on both sides of the ocean have similar subjects to deal with, meet with the same difficulties, expose themselves sometimes to similar dangers; they are like companions in arms in the battle fought by mankind in order to subjugate the natural forces and to draw profit therefrom. There is no other field where this fraternity must manifest itself more actively than in the studies pursued with a view to improving the safety of mines and to rendering impossible in the future these mining disasters, which every year claim hundreds of victims. These are very difficult problems, the solution of which requires numerous and laborious experiments accompanied by very arduous scientific investigations; it is extremely desirable that the various mining countries, which have undertaken to solve these difficult problems, should unite their efforts in order to succeed as soon as possible and as well as possible. This idea has been fully grasped by your Bureau of Mines, which has brought me to Pittsburg; the daily exchange of views with the distinguished chief of-the mine investigation division, George S. Rice, and with his excellent collaborators, the comparison of the results obtained at Lievin and Bruceton, the checking of these results by means of new tests, the profound study of the measurements, diagrams, measuring and laboratory apparatus, the cooperative investigation of new apparatus, or the preparation of new investigations, are, I hope, of such a nature as to bring about great progress. I think that it will be interesting to you to know what has been done in France along this line.

Jan 9, 1914

The ore milled from the Lead Lode of The Zinc Corporation leases consists of a fairly granular mixture of galena, marmatite, chalcopyrite, tetrahedrite, pyrrhotite, pyrite, and lollingite with a gangue association of calcite, quartz, bustamite, rhodonite, mangan-hedenbergite, garnet, fluorite, mangal1o-calcite and apatite. The silver content of the ore occurs in three minerals-tetrahedrite, dyscrasite, and pyrargyrite, the last two mentioned occurring in microscopic association only. Other metallic and gangue minerals are also present in this complex mineralogical association, but do not influence the milling operations. Fortunately, the silver minerals favour the galena and chalcopyrite as their host minerals, thus facilitating the recovery of both copper and silver in the lead concentrate. The ore from the Lead Lode assays from 14-18% lead; 3·0-4·5 oz. silver per ton, and 9·5-12·0% zinc.A parallel lode, known as the Zinc Lode, is well developed in The Zinc Corporation leases. This lode assays from 6-8% lead, 1·5 oz. silver per ton, 18-25% zinc with a siliceous gangue. The common Lead Lode gangue minerals-ealcite, rhodonite, bustamite-are very rare in present underground exposures. No metallurgical work was done on this ore during the pilot plant operations.The aim in the milling process is to produce: 1. A lead concentrate with the maximum economic recovery of lead, silver, gold, copper and antimony, and minimum zinc content.2. A zinc conc,entrate with the maximum economic recovery of zinc, cadmium, and cobalt, and minimum iron, arsenic and fluorine content.3. A waste residue containing sufficient granular material to be utilized as underground fill.In considering the economics of the milling practice the above three requirements plus consideration of the tariffs submitted by the purchasers of concentrates were the deciding factors in the evolution of the present milling practice.Towards the end of 1924 The Zinc Corporation Limited added to and adapted the old zinc tailings plant for the treatment of 5,000 tons of crude ore per week. This plant, which operated until January, 1936, was in general principles similar to the standard Broken Hill methods of ore concentration, whereby the galena was recovered by the gravity concentration methods of Hgging and tabling with selective flotation concentration of the residual galena and marmatite. One important difference from other Broken Hill plants was that the primary grinding of the jig tailing was performed with flint-loaded mills operating in closed circuit with revolving screens. Two primary grinding sections were installed, each...

Jan 1, 1941

By YOUNIS S. MUSTAFA, GHORASHI ZAIN, ABDEL-ZAHER M. ABOUZEID, AHMED A. S. SEIFELNASSR

The present paper attempts to characterize the breakage properties of a brittle rock, quartz, at different size fractions when compressed at various loads in a confined piston die.  This is to understand the interaction between the various parameters affecting the breakage behavior of a brittle solid in a confined zone. Application of the results may lead to improving the energy efficiency in the size-reduction units, which use compression as the main stressing force. 

By Bo Jiang, Jiusan Xiao, Hongmin Zhu, Kai Huang, Shuqiang Jiao

Carbothermal selective reduction of simulative titanium slag and various titanium raw materials were presented and SiO2 was found to be partially reduced, forming a series of Fe-Si alloys during the process. Thermodynamic analysis of the selective reduction between TiO2 and SiO2 was performed using HSC 6.0 software, then a series of equilibrium experiments were designed and presented. The activity of Si reduced from SiO2 was affected by the presence of Fe reduced from FeOx and the generation of Fe-Si alloys, promoting the reduction of SiO2 in the whole process. The results of experiment are in great accordance with the calculated data obtained from thermodynamic analysis.

Jan 1, 2016

By James J. Meadows, Jason D. Pampena, Emanuele G. Cauda, Lauren G. Chubb

Respirable crystalline silica (RCS) found in dust generated by various mining processes is an important occupational hazard in the underground coal-mining industry. Exposure can lead to disabling and incurable diseases, including coal workers’ pneumoconiosis (CWP), silicosis and lung cancer [1,2]. Clusters of rapidly progressive CWP have been observed in the Appalachian regions of eastern Kentucky and southwestern Virginia, and recent cases have been generally more severe and showing more rapid progression. CWP is approximately four times as common among long-tenured underground coal miners in central Appalachia than long-tenured underground miners in the rest of the United States, and 20.6 percent of long-tenured underground miners in central Appalachia were suffering from the disease as of 2017.

By Douglas Bunting

(Wilkes-Barre Meeting, June, 1911.) WITH the gradual exhaustion of the upper veins in the anthracite coal-fields, the problem of mining at greater depths acquires increasing importance and demands the consideration of a number of important factors, one of which is the greater earth-pressure and the consequent necessity of stronger support o for the roof. Under the pillar-and-chamber system, almost exclusively followed in the Northern anthracite-field, the highest economy in mining is generally secured by leaving, on first mining, pillars only sufficient to support safely the overlying strata. As to the necessary size of such pillars, the opinions of mining experts are widely divergent. In establishing the width of chambers and pillars, the thickness of vein and its depth below the surface have received little consideration. For. instance, it has been quite usual to work both overlying and underlying veins with the same width of chambers and pillars, when the lower vein was two and one-half times as thick, and twice as fits below the surface, as the upper. In view of the generally-accepted theory that the crushing-strength of coal-pillars of the same base-area becomes less with increased height, it is probable, in this instance at least, that the most economical mining has not been secured. This question of adequate pillar-support is economically less important down to about 800 ft. than at greater depths; and it is with reference to mining at these greater depths that the study of the subject here presented has been prompted. The necessity of leaving larger pillars, involving greater mining-costs per ton and also smaller yields per acre, is one of the troubles of deep working. To mine without leaving adequate pillar-supports will result, sooner or later, in a squeeze. Limited areas, it is true, have been mined at certain widths of

Sep 1, 1911

By W. H. Fluker

Until recently, the universal practice in New Zealand was dry-crushing and direct cyaniding. With ores containing no mineral sulphides, and little or no coarse gold, 'this method, in spite of its many disadvantages, gave excellent results, and was by far the best one available, pending the solution of the slimes-problem. But the situation has now been changed. The successful treatment of the slimes, enabling the mines not only to do away with dry-crushing, but also to deal with the mineralized ores of the lower levels, which are, for obvious reasons, quite unsuited to that method of treatment, has made wetcrushing universal; and this may be taken as satisfactory proof of its superiority over the former practice. In the lack of accurate information from plants operating successfully, particulars of experimental trials carried out on a working-scale at the Waitekauri Extended mine, Maratoto,

Jan 1, 1903

By A. J. van der Zeeuw

The use of SME 529 in the extraction of copper from acidic liquors and of copper and nickel from ammoniacal solutions, is illustrated with results of laboratory and bench scale mixer/settler experiments. Special features of SME 529 are its high loading capacity and the ability to keep the complexes dissolved in the organic phase up to quite high concentrations.

Jan 1, 1976

By D. R. Tesarik

To ensure the stability ofa tailings pond embankment, the height of the phreatic surface must be kept at or below a safe level. In this investigation, the Bureau of Mines analyzed various horizontal drain designs for tailings embankments to determine their effects on location of the phreatic surface. This report describes the investigation, and it includes 21 dimensionless graphs that can be used to estimate the drain spacing and length dimensions necessary to ensure the stability of embankments of various configurations. Analyses were based on the use of two computer codes, a three-dimensional finite-element code and a two-dimensional finite-difference code. The computer-generated results were compared with results obtained from a laboratory embankment model, other laboratory test results, and piezomet-ric data from two actual tailings embankments. Nearly the same phreatic surface locations were predicted using either of the computer codes. For one of the actual embankments studied, the predicted phreatic surface location was slightly higher than the measured location; for the other, the predicted location closely followed the actual surface trend. Phreatic surfaces of the laboratory model were slightly higher than the code-gener-ated locations, and the differences grew larger as drain length increased or drain spacing decreased.

Jan 1, 1984