Search Documents

By C. W. Vierrether

Mineralogical analysis contributes directly to the selection of applicable beneficiation techniques for a given site. Tentative remediation methods for use on the Sharon Steel tailings in Midvale, Salt lake County, UT, were determined in part by mineralogical characterization. A systematic characterization scheme was employed to identify mineral species, mineral associations, and grain size. This paper presents the mineral content and distribution data collected for site characterization. This data was used to determine areas that were representative of the total tailings at the Sharon Steel site. Although this was the main focus of characterization, other useful information was obtained. The approximate location of "old jig tails" was identified and insight on possible arsenic, mobility through the pile was also obtained from the mineralogical study.

Jan 1, 1991

By Homor N. McDonald, G. Dale Coe, Will H. Coghill

"INTRODUCTION Sporadic beneficiation tests have been made on the Birmingham (Ala.) red iron ores for the last 55 years. Now, gravity-concentration plants are in operation, of tables are in use. A major problem, as seen by the present writers, is to obtain a coarser concentrate and add to the recovery without increasing the quantity of slime. The oily practicable solution found so far is to recover, by means of concentration (hydraulic) classifiers, the coarse concentrate normally lost in the table middling and, by a similar method, to recover middling particles from the table tailings The size and shape of the coarse concentrate Particles combine to cause them to report in the table middling rather than in the table concentrate. Similarly, mineral is lost in the tailing. Because of this, the nature of the ore, the deportment jerking table, plant tests on classifiers and tables, and the suitability of small concentration classifiers to treat table zonal middlings are discussed herein.The work was done by the Southern Experiment Station, Bureau of Mines, in cooperation with the University of Alabama, Tuscaloosa, Ala."

Jan 1, 1935

By Jarrett. K. M.

Nobody has, as yet, made a perfect machine. Therefore, when a machine is operating and functioning as it is designed to do, there is always some vibration present. This is easily and simply measurable with present-day equipment and any increase in the amount of vibration will indicate a potential failure. When an increase is noted, vibration analysis can pinpoint the cause of the potential failure and remedial action can be taken before an unscheduled shutdown is necessary or a breakdown occurs.

Jan 1, 1972

By R. F. Hochstein, A. J. Neale, B. C. Flintoff

"Process control technology holds very significant potential benefits for many mineral and coal processing plants. The impact can be expected in terms of performance improvements (e.g. increased throughput, better quality control, higher availability), through reduced operating costs (e.g. lower reagent consumptions, manpower reductions) and in the form of intangibles (e.g. reduced vulnerability to manpower turnover, increased productivity in the user group). However, the installation of a process control system will represent considerable capital expenditure and will generate an operating cost. No management team would approve the funding for such a project without some reasonable measure(s) of the profitability of a proposed system. This paper describes a methodology for addressing the justification problem. Of particular interest is the approach to quantifying performance improvements, frequently found to be the predominant benefit of process control.INTROPUCTIONProcess control has become widely recognized as an essential component of any processing operation. It has direct impact on productivity, process efficiency and product marketability. The reported pay backs associated with successful installations underscore this observation. As a consequence, many operators who have little in the way of process automation are looking to initiate control projects, while those who currently enjoy some level of automation are looking to upgrade existing control systems (= process + instrumentation + hardware + software + operators + support). While the emphasis here would appear to be on retrofit process control projects, new plants, or those currently in the design stage, will usually find themselves in one of these categories shortly after startup 1."

Jan 1, 1990

Since 1955 Gaspe Mines has produced 144 million tonnes of ore, averaging 0.87% Cu, mined from two open pits and underground workings. In 1994, four exploration drill holes discovered a buried porphyry/skarn Cu-Mo deposit at depths of between 1000 m and 1700 m below the top of Porphyry Mountain. The mineralization and alteration associated with the Porphyry mountain quartz monzonite does not outcrop. A total of 19,500 m of drilling in 12 drill holes has outlined 200 million tonnes grading 0.73% Cu and 0.08% Mo. Mineralization occurs as veins and disseminations of chalcopyrite and molybdenite. Drilling and feasibility studies continue.

Jan 1, 1998

By Gloyd M. Wiles

The property of the Park City Consolidated Mines Co. is situated in the northeast part of the Park City dining district 1 mile east of the town of Park City. The elevation at the collar of No. 1 shaft is 7,250 feet. Park City is a town of approximately 4,500 population lying about 30 miles east of Salt Lake City. The two are connected by an oiled State Highway, which is cleared of snow in winter and maintained in first-class con¬dition by the State throughout the year. The district is served by branches of the Denver & Rio Grande Western and Union Pacific Railroads, each operating on a daily' schedule, except Sunday. These two railroads have constructed 1 ½ miles of joint track from the town to the Park City Consolidated Mines Co. loading tipple, located 200 feet from the portal of the adit that leads to No. 1 shaft.

Jan 1, 1936

By Chad W. Brown

"A study of five commercially available solvent-extraction diluents in combination with three commercially available extractants under sub-Saharan African copper solvent-extraction conditions yielded intriguing results. A variety of physical and chemical property analyses showed that the diluents have some property differences, but overall the properties are similar and provide comparable solvent-extraction performance. Although each physical chemical property technique provides useful information, no individual test result suffices for prediction of diluent performance. Furthermore, a diluent should be evaluated within the conditions for which it is targeted to determine a true measure of performance. Overall, the diluents studied possess appropriate physicochemical properties for sub-Sahara Africa copper solvent-extraction conditions and can be utilized to achieve efficient, safe and reliable operations. INTRODUCTION Background Today, commercial SX diluents are available globally from a limited number of manufacturers. Production quality and consistency of these diluents have improved tremendously over the past 20 years. For example, diluent flash points have been increased incrementally to improve transport and handling while reducing fire hazards. In addition, properties have been successfully modified to minimize evaporative losses, ultimately reducing SX plant raw material costs. Diluent manufacturers have also focused on meeting performance requirements of SX plants, including eliminating components detrimental to SX operations, such as olefinic groups that can react with various species resulting in changing phase-disengagement properties over time, as well as ensuring that proper hydrocarbon components are present within diluent mixtures to maintain good reagent–metal complex solubility. These challenges have been met in the midst of ever-increasing stringent environmental regulations. Sub-Saharan Africa copper SX operations pose a unique set of conditions for diluents due to inherent ore compositions and the associated high reagent concentrations required for complete copper extraction. Chevron Phillips Chemical Company, LP has been manufacturing hydrocarbons and hydrocarbon-based diluents since the mid-80s, applying this experience to identify the physical properties considered crucial for proper performance in SX operations. This paper documents the findings of a study utilizing commercially available diluents in combination with extraction reagents commonly utilized in African SX operations to evaluate the relative performance of the diluents under sub-Saharan Africa copper SX conditions."

Jan 1, 2015

By Ronald Gibala

THE Au-Ni alloy system is one of many binary systems in which the supersaturated solid solutions may exhibit a multistaged precipitation behavior during aging. In addition to the equilibrium phase separation, which was studied extensively years ago in the range 400" to 600",' there exists at least one earlier stage of precipitation, i.e., preprecipitation, which occurs below 300°C in quenched alls.-This note reports on some effects of composition on the preprecipitation in Au-Ni alloys as determined from resistivity measurements. The kinetics of the preprecipitation were studied in alloys of 15 to 85 at. pct Ni which were water- were found in alloys which after various quenching-and-aging treatments were slowly heated and cooled (1Z"C per hr) between room temperature and about 300°C. Two such heating-cooling curves for a 40 pct Ni alloy are shown in Fig. 2. Similar effects were observed in other 15 to 50 pct Ni alloys. These curves are characterized upon heating by two changes in slope indicated by A and B in the figure and upon cooling by a resistivity hysteresis of approximately the size of the original decrease in resistivity after the aging treatment. The inflection at B generally falls at about the same temperature for all heating curves for a given alloy and appears to be a critical temperature (T,) for the preprecipitation for that alloy. The values of T, for most of the 15 to 50 pct Ni alloys fall in the range 200" to 220°C. The position of the inflection at A depends upon the heat treatment prior to the heating-cooling run and perhaps the heating rate, which was held constant in these runs. The preprecipitates grown in by the appropriate aging treatments seem to decrease in size as the alloys are heated from A to T,, dissolve at T,, and do not reappear after cooling below T,. When similar heating-cooling treatments were performed on 60 to 85 pct Ni alloys, an entirely different behavior was observed. Typical results are shown in Fig. 3 for an 85 pct Ni alloy. In these

Jan 1, 1964

The discovery of gold mineralization at the Manhattan mercury mine, Napa and Yolo counties, California in 1978 was the result of applying a new exploration ore deposit model in a region previously unprospected for gold. This discovery, the McLaughlin gold mine, owned and operated by Home-stake Mining Company, is a classic epithermal mercury-hot springs-gold deposit currently producing approximately 7,776,000 grams of gold annually.The exploration program that led to this discovery was based on a mercury-hot springs-gold genetic model. The model was developed using basic knowledge of ore-forming processes plus careful field observations made during examination of many known precious metal districts in. the Western U.S.A. Theoretically, hot spring systems could host gold deposits and, indeed, gold in these environments had been recognized for years. However, at the time of the McLaughlin discovery, it was not known to occur in sufficient quantities to inspire geologists to explore hot spring systems specifically for gold. Field observations recognized the presence of hot spring deposits and mercury mineralization asso-ciated with precious metal districts, (Bodie, California being a type example), and this associ-ation prompted the development of a hypothetical model. Subsequent exploration activities were directed towards the examination of mercury mines and prospects as potential targets for gold deposits. The model envisioned hot spring deposits (sinter) containing mercury mineralization at the surface with anomalous gold values grading downward into significant ore grade gold values.

Jan 1, 1990

By Nabil F. Ismael, Khalid Farraj

"The axial load transfer along large diameter bored piles in very dense silty sand was examined by field tests. Two instrumented piles 1.2 m diameter and with embedded length of 18 m and 25 m were load tested in axial compression to failure in connection with the construction of a big tower building in Kuwait City, Kuwait. Instrumentation included vibrating wire strain gages at various levels, displacement transducers and extensometers. The purpose being to measure the axial load distribution, skin friction, and base resistance at various load levels. In addition the settlement of the piles and the elastic compression were monitored. The ultimate pile capacity, side friction, and base resistance were determined and correlated with the Standard Penetration Resistance values as determined from the boreholes. Most of the load was transferred as side friction and the pile settlements at the maximum loads ranged from 1.6% to 2.2% of the pile diameter.INTRODUCTIONLarge diameter bored piles are used with increasing frequency for heavy structures in Kuwait and the Gulf region in calcareous sand. The behavior of short piles ranging in depth from 8 m to 15 m was examined by field test (Ismael 2012). However, larger diameter and deeper piles are being used more frequently, including the project described in this paper which included piles with 1.2 m diameter and 18 m to 25 m depth. The design of these piles requires knowledge of the skin friction and point resistance in these soils. While field tests on driven piles are available (Ismael 1989, Ismael 1999), field test data on long large diameter piles installed through very dense sands or silty sands are very limited at present. The results of two load tests carried out recently on instrumented bored piles at one site in downtown Kuwait City have become available. In these tests the failure load was extrapolated from the test results. It is important to examine and analyze the test results for the benefit of the geotechnical research and design engineers."

Jan 1, 2017

By W. H. Brandt

ThEre has been much progress in the last two decades in understanding the hardenability of steel. Roughly, the progress has been along two lines, which may be designated as empirical and fundamental. This empirical work has progressed to the point where steels may be manufactured to the desired characteristics quite closely and heat-treatment and application problems can be approached with considerable confidcnce. Particularly noteworthy among the empirical work is that of Grossmann,12. who has given formulas for calculating hardenability as a function of composition, and the establishment of the isothermal transformation curve by Davenport and Bain.13 The latter does not fall entirely in the class of an empirical study, for while it provides a means of presenting a tremendous amount of empirical data in the form of a single curve, it also outlines clearly the problems that face the fundamental investigator. The isothermal transformation curve shows clearly that the phenomena occurring in the neighborhood of the knee of the curve are those that limit the hardenability of steel. Above the knee of the curve, the product of decomposition of austenite is lamellar pearlite and fundamental progress has consisted chiefly of a detailed study of structure and formation of pearlite, largely at the hands of Mehl and his collaborators.'-5 They have shown that formation of pearlite proceeds by a process of nuclea-tion and growth and they measured both the nucleation and growth rate. However, there is still a wide gap between our empirical knowledge and fundamental knowledge, owing largely to the fact that there is no satisfactory theory of nucleation rate or of growth rate. This paper is being written in an attempt to Hl in a small part of the gap. A method was prcsented recently1' for calculating the velocity of edgewise growth of pearlite using an expression for the carbon concentration that satisfies the diffusion equation. The edgewise velocity of pearlite growth in a plain carbon eu-tectoid steel was calculated, using linear extrapolations of the ferrite and comentite solubilities. While the results are in fairly good agreement with the experimental facts, the calculated velocity of edgewise growth is smaller at 600°C. than G, the radial growth velocity of pearlite nodules, whereas one would expect it to be larger. In this paper the calculation will be repeated using a new extrapolation for the ferrite solubility, which gives better agreement between G and T', the calculated velocity of edgewise growth. A calculation will also be made of the effect of carbon content on velocity of edgewise growth, and the problems of calculating the effects of other alloying elements on growth rates will be discussed. It will be assumed that the mechanism of pearlite growth, established by Mehl and

Jan 1, 1947

By W. H. Brandt

ThEre has been much progress in the last two decades in understanding the hardenability of steel. Roughly, the progress has been along two lines, which may be designated as empirical and fundamental. This empirical work has progressed to the point where steels may be manufactured to the desired characteristics quite closely and heat-treatment and application problems can be approached with considerable confidcnce. Particularly noteworthy among the empirical work is that of Grossmann,12. who has given formulas for calculating hardenability as a function of composition, and the establishment of the isothermal transformation curve by Davenport and Bain.13 The latter does not fall entirely in the class of an empirical study, for while it provides a means of presenting a tremendous amount of empirical data in the form of a single curve, it also outlines clearly the problems that face the fundamental investigator. The isothermal transformation curve shows clearly that the phenomena occurring in the neighborhood of the knee of the curve are those that limit the hardenability of steel. Above the knee of the curve, the product of decomposition of austenite is lamellar pearlite and fundamental progress has consisted chiefly of a detailed study of structure and formation of pearlite, largely at the hands of Mehl and his collaborators.'-5 They have shown that formation of pearlite proceeds by a process of nuclea-tion and growth and they measured both the nucleation and growth rate. However, there is still a wide gap between our empirical knowledge and fundamental knowledge, owing largely to the fact that there is no satisfactory theory of nucleation rate or of growth rate. This paper is being written in an attempt to Hl in a small part of the gap. A method was prcsented recently1' for calculating the velocity of edgewise growth of pearlite using an expression for the carbon concentration that satisfies the diffusion equation. The edgewise velocity of pearlite growth in a plain carbon eu-tectoid steel was calculated, using linear extrapolations of the ferrite and comentite solubilities. While the results are in fairly good agreement with the experimental facts, the calculated velocity of edgewise growth is smaller at 600°C. than G, the radial growth velocity of pearlite nodules, whereas one would expect it to be larger. In this paper the calculation will be repeated using a new extrapolation for the ferrite solubility, which gives better agreement between G and T', the calculated velocity of edgewise growth. A calculation will also be made of the effect of carbon content on velocity of edgewise growth, and the problems of calculating the effects of other alloying elements on growth rates will be discussed. It will be assumed that the mechanism of pearlite growth, established by Mehl and

Jan 1, 1947

By George White

On June 25, 1876 Lt. Colonel George Armstrong Custer led the Seventh U.S. Cavalry Regiment., comprising 31 officers and 637 enlisted men, into the valley of the Little Big Horn River in southern Montana. His assigned mission was a reconnaissance-in- force to locate the Indians reported to be assemblying in that area. Custer was extremely successful in locating the Indians. Unfortunately he located more than 3000 Cheyenne and Sioux under the leadership of the great war chiefs Crazy Horse and Gall. Custer's entire battalion of more than 200 men was annihilated; there were no survivors. While I don't seriously suggest that things will be all that bad in the uranium industry, I do suggest that some very fundamental changes have occurred in the uranium market. Conventional thinking based on past truths will no longer suffice. Geology, economics, politics, nuclear growth pat terns, technology, and utility attitudes have all changed. New thinking is in order. It is at this point that I would offer my usual disclaimer. My views regarding the future are simply that--my personal views. My belief that certain events will come to pass does not mean I view those events as appropriate, wise, good, equitable, or economic.

Jan 1, 1985

By Raymond Mancha

TO determine the optimum inside dimension of an airshaft, it is necessary to strike the proper balance between the cost of power for air friction and turbulence losses within the airshaft, on the one hand, and the cost of sinking the airshaft, on the other. An airshaft larger than necessary is extravagant because the increased amortization rate exceeds the reduction in power cost. Conversely, an airshaft smaller than is needed wastes more power than can be justified by the corresponding reduction in amortization rate. The airshaft power is a function of the air volume Q cu. ft. per min. and the air-pressure drop p in. water gauge across the airshaft. † The airshaft pressure drop is calculated by the formula: R X l X O X Q2 P= 5-2 X A3 wherein R is the air-friction coefficient for the portion of the airshaft 1 ft. deep, 0 ft. in perimeter and A sq. ft. in area, when passing an air volume of Q cu. ft. per minute. The airpower w kw. dissipated within the airshaft is computed by the formula: w= 0.746Q X P 6345 With a combined over-all unit efficiency of the fan and drive of E per cent, the air-shaft power requirements become W = 100w There are 8760 hours per year, so with energy at I cent per kilowatt-hour the annual cost per kilowatt for continuous operation is $87.60; therefore, with energy at C¢ per kw-hr. the annual airshaft power cost MI dollars per year becomes: M1 = 87.60W X C which, expressed by the component parts, becomes: M1 V 5.2X4' / XQ Z 0.746 X 100 X 87.60 X C V 60 X E ) The amortization rate a dollars per year per dollar invested includes both depreciation and interest on the investment. These in turn involve the expected useful life of the airshaft and available molletary interest rate. For the purpose of this discussion the cost of airshaft sinking is assumed to comprise a fixed move-in charge Y dollars, which covers the cost of transporting, erecting and removing the shaft-sinking equipment and is independent of the air-shaft size. Also, it is assumed that the cost of shaft lining, if any, is fixed at $X per cu. yd. of lining material and that excava-

Jan 1, 1947

By Raymond Mancha

TO determine the optimum inside dimension of an airshaft, it is necessary to strike the proper balance between the cost of power for air friction and turbulence losses within the airshaft, on the one hand, and the cost of sinking the airshaft, on the other. An airshaft larger than necessary is extravagant because the increased amortization rate exceeds the reduction in power cost. Conversely, an airshaft smaller than is needed wastes more power than can be justified by the corresponding reduction in amortization rate. The airshaft power is a function of the air volume Q cu. ft. per min. and the air-pressure drop p in. water gauge across the airshaft. † The airshaft pressure drop is calculated by the formula: R X l X O X Q2 P= 5-2 X A3 wherein R is the air-friction coefficient for the portion of the airshaft 1 ft. deep, 0 ft. in perimeter and A sq. ft. in area, when passing an air volume of Q cu. ft. per minute. The airpower w kw. dissipated within the airshaft is computed by the formula: w= 0.746Q X P 6345 With a combined over-all unit efficiency of the fan and drive of E per cent, the air-shaft power requirements become W = 100w There are 8760 hours per year, so with energy at I cent per kilowatt-hour the annual cost per kilowatt for continuous operation is $87.60; therefore, with energy at C¢ per kw-hr. the annual airshaft power cost MI dollars per year becomes: M1 = 87.60W X C which, expressed by the component parts, becomes: M1 V 5.2X4' / XQ Z 0.746 X 100 X 87.60 X C V 60 X E ) The amortization rate a dollars per year per dollar invested includes both depreciation and interest on the investment. These in turn involve the expected useful life of the airshaft and available molletary interest rate. For the purpose of this discussion the cost of airshaft sinking is assumed to comprise a fixed move-in charge Y dollars, which covers the cost of transporting, erecting and removing the shaft-sinking equipment and is independent of the air-shaft size. Also, it is assumed that the cost of shaft lining, if any, is fixed at $X per cu. yd. of lining material and that excava-

Jan 1, 1947

By J. M. Ortiz, J. Haas, W. Kracht, G. Pamparana

This work explores the effect of ore hardness variability on the design of a solar photovoltaic system with battery storage and grid backup, to operate a semi-autogenous grinding mill (SAG). First, a case of two synthetic ore populations was studied, without any consideration of spatial distribution. The mean and variance of the ore hardness (measured in specific energy consumption, SEC) of both populations were subjected to a sensitivity analysis. The operation was evaluated with one and two stockpiles. The latter allows classifying according to ore hardness as a way to adapt the SAG energy consumption to the solar cycle. The results show that using two stockpiles allows achieving a larger autonomy (i.e. less energy imports) at lower total costs, especially when the difference of mean hardness between the two populations is large. The difference in variance (between the two populations) has a smaller impact on the design recommendations. Second, a case study was performed considering the spatial distribution of geometallurgical units. Two cases were considered: (a) brownfield, when there is historical information of the actual variability of the ore hardness; and (b) greenfield, when the system design relies mainly on geometallurgical modeling. Brownfield and greenfield planning result in similar sizing recommendations of the solar power plant, if precise drill cores are available. In this case, geometallurgical modeling is an effective tool for forecasting the energy demand necessary for designing the power plant.

Jan 1, 2018

By C. W. Bollier

In recent years, Peabody Coal Company has enjoyed a great amount of success using roof trussing as a secondary support system. Since 1975, over 180,000 trusses have been installed in 12 different Peabody mines with varying geological conditions. Projections from Peabody, and throughout the mining industry, indicate a continuous growth in the trussing principle for roof control. Peabody's primary supplier has sold over 1,000,000 trusses to 50 different mines throughout the country. (1) Roof trusses provide an improved means of controlling difficult roof conditions which, in turn, improves worker safety and mine production. Since the original development of the truss, (2) the only major modification to the trussing procedure has been the addition of the resin point anchor system in the mid-70's. The resin point anchor was the catalyst needed to provide a reliable anchor and create acceptance of the roof truss by both management and labor. Peabody began using trusses in their mines in 1975. For the first few years, most of the truss effort was geared toward determining proper anchorage, spacing, and gathering historical data. In recent years, Peabody's activities have centered on the development of improving trussing procedures. Through monitoring and trussing activities, a number of areas of improvement have been found.

Jan 1, 1982

By A. C. Dittrick

THE Simon-Carves washer, originally known as the Baum jig washer, was first built in Europe about 1892. It was probably the first washer to use intermittent pressure of air on water in one part of the wash box to obtain water pulsations in another part. This type of washer was also the first to successfully clean various sizes in the same unit. Previous practice was to size closely before washing. Today there are several large firms abroad, and a few in the United States, that are constructing the Baum type washer. The design of all is practically the same, their differences being only in details. In 1928, the Hanna Coal Co. first gave consideration to the question of coal cleaning. It was believed at that time that by efficient removal of so-called noncombustibles appearing in the Pittsburgh No. 8 seam, certain markets that had been lost due to poor preparation might be recovered. Mechanical cleaning and mechanical loading were under consideration. This type of cleaning was considered the most practical means of removing impurities in the sizes below 4 in. It was decided to adopt mechanical cleaning and develop a system of mechanical loading that would be successful on No. 8 coal.

Jan 1, 1935

By D. Tahija

"Carbonic acid prepared on-site from carbon dioxide gas is widely used in water treatment and increasingly in some industrial applications. It has not seen extensive use in mineral processing but is an attractive alternative to sulphuric acid for pH control, with the need to counteract high pH levels caused by cemented backfill of particular interest. The fundamental chemistry of carbonate solutions as applied to flotation and flocculation circuits is described as well as plant-scale application. A project to introduce use of carbon dioxide for pH control at the Hecla Greens Creek lead-zinc-silver-gold operation is planned to begin operation in the first half of 2015. INTRODUCTION Flotation pH Control and Backfill Most sulphide flotation circuits are run at pH values ranging from slightly acid to strongly basic, with slightly basic pH the most common. Lead sulphide circuits are typically run at values from 7.5 to 9, with zinc circuits running at higher values of 9.5 to 10.5. This has been the case with ores that produce a natural pH in the circumneutral range, so that the pH needs to be raised. Although lime [Ca(OH)2] is the cheapest reagent in use to raise pH, soda ash (Na2CO3) or caustic soda (NaOH) are also employed where sodium ions in solution are preferred to dissolved calcium, or, in the case of soda ash, the presence of carbonate is desired. The use of soda ash for pH control in lead circuits has been widely recommended for many decades (1). The situation is necessarily different at operations where cemented backfill is used in underground mines (2). Backfill, whether emplaced as rammed paste or pumped slurry, is composed of mill tailings mixed with around 10% Portland or similar cement. When later mining occurs near or adjacent to previously backfilled areas, a considerable amount of old backfill is inevitably mixed with new ore and shipped to the mill. Depending on orebody geometry, mining methods etc., the amount of backfill in mill feed can easily approach 20%, with periods of higher content not unheard of. The presence of so much backfill, with its high cement content, produces results largely indistinguishable from adding a large amount of lime to run-of-mine feed: mill pH in the grinding circuits can range above 11 and high concentrations of calcium ions are present. The need to raise pH disappears; the problem becomes to lower it."

Jan 1, 2015

By Dr. David R. Hargis

Finite-difference digital computer models can be used to evaluate the effects of proposed mine dewatering and depressuring operations on ground-water aquifer systems. The effects of proposed dewatering and depressuring operations can be simulated for both underground and open-pit mining in confined and unconfined aquifer systems. Design of dewatering and depressuring well arrays to accommodate proposed mining operations in a variety of hydrogeologic settings can be accomplished through judicious application of well-documented and tested computer programs that are widely available from government agencies and universities. The speed and flexibility of the computer models can provide an economical means for comparison of alternative mining plans, and can be used to determine the optimum mine plan with respect to dewatering requirements.

Jan 1, 1980