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By John Bailey

Stressmap is a program whereby the earth's stress field is charted in detail and utilized for design of mines and mine exploitation. This program is implemented by mapping the attitude (location, direction, dip, and extent) of induced rock ruptures during the process of hydrafracturing drill holes contiguous to mine interest. The program includes utilization of ancillary naturally occurring stress relief signals, They are precipitated because the earth is indeed dynamic and the strata of the crust are continually adjusting to load and/or strain differentials resulting in rock deformation and fracturing. These fracturing adjustments occur as small to tiny earthquakes comparable to Richter magnitudes of one, zero, minus one, and smaller events termed microearthquakes. Obviously, the earth's stress field changes due to continued differential loading and crustal movements; therefore it becomes conspicuously evident that resolving today's stress field as opposed to paleostresses is most important for efficient, economical, and safe mining operations.

Jan 1, 1974

By J. Rostami, P. P. Nelson, A. Khetwaln

Tunnel boring machines (TBMs) can be considered as the method of choice for tunneling in many projects due to the high level of efficiency and high potential productivity, safety, and applicability in a variety of subsurface conditions. The muck transportation system in a TBM tunneling operation plays a significant role in the process and impacts the overall performance of the machine. This paper introduces a recently developed method of simulating TBM operation through discrete event simulation (DES), which allows for tracking of the various activities, incurred downtimes in the operations, and overall machine utilization. The study presented here compares two different muck transportation systems, conveyor and rail, under similar geological conditions to quantify the impacts of the choice of muck transportation systems on machine utilization. The results show the applicability and extent of redundancy in the use of the two systems relative to machine capabilities and their sensitivity to geological conditions.

Jun 13, 2021

By S. E. Phifer

The Haile Gold Mine was the largest gold producer east of the Mississippi River prior to 1942. It is currently the only producing gold mine in the Southeastern United States. Current development involves the mining and heap leaching of some 1000 TPD. Carbon is employed as a recovery medium. Due to its clayey nature and excessive fines in the oxide ore, agglomeration using Portland cement is required prior to leaching. The ore ranges in screen size from two inches to -200 mesh. The agglomerating step combined with certain other conditions has enabled Piedmont Mining Company to achieve recoveries of 90%.

Jan 1, 1987

By D. J. Beeby

California's Surface Mining and Reclamation Act (SMARA) was passed in 1975 to identify and protect mineral resources in areas of high land use conflict and ensure reclamation of mined lands. Under SMARA, aggregate resource availability is quantified and compared to forecasted demand in specific market regions. Fifteen regions studied thus far cover 51,800 km2 (20,000 sq miles) and contain 45 Gt (50 billion st) of high-quality aggregate resources that have been legally designated as "regionally significant," providing a measure of protection previously unavailable. SMARA reports and state testimony at local hearings have supplied objective information and helped reduce conflict in mine permitting decisions.

Jan 1, 1987

By D. A. Wheeler

Invented by Pierre Boutin in the early 1960s the column was a complete break from the conventional flotation cell. 1. When used as a rougher-scavenger, the column is excellent. 2. When used as a cleaner, the results can be spectacular. The very first column tests were carried out using it as a rougher-scavenger in a reverse float where silica was floated from iron. It produced a concentrate underflow as good or better than that produced from roughing and scavenging stages in cells, The froth tailing overflow was better than that produced by several stages of cell cleaning. Column scaleup progressed rapidly from the two inch diameter unit to a semi automated 12 inch diameter column on material from the Iron Ore Company of Canada. A change in operating philosophy within IOCC brought all flotation development, column or conventional, to a halt. At that time, IOCC had an exclusive right to the use of the column in Canada in the field of iron ore. We moved into the field of sulphides. A Canadian copper producer sent ore for testing and the results led to their purchase of the first commercial size of column - a 36 inch diameter machine. It was a mechanical disaster. It took several years to raise sufficient funds to return to that mill with our basic 18 inch square unit. It was to be tested and modified in order to learn how to properly design a large column. Originally used as a rougher- scavenger, it had to produce tails equal to the final tailings from this well run plant. It did, and did so while producing a rougher-scavenger concentrate almost equal to the plant final concentrate. It was finally used as a cleaner and produced concentrate 5% higher in copper than the plant final concentrate with equal cleaner tails. This phase of the column development was carried out under very difficult circumstances. The mill superintendent had realized that if he took the froth removal system from our original mechanical disaster and applied it to a conditioner while injecting air, he would have a flotation cell. He had personally applied for patents on this Maxwell Cell. Our development work was done in his mill and as our results became better and better, our difficulties became worse and worse. We finally had to terminate this work. The 18 inch column installed at Mines Gaspé 14 years later was identical to the one removed from Opemiska. The 18 inch column was tried on various ores over the following years and always produced excellent results. However, it was not really a production size unit. We had always aimed for the 72 inch column (72" x 72" x 44' 9"). Prior to this huge machine, we needed the intermediate 36 inch column. The failure of the original 36 inch diameter unit at Opemiska had raised the possibility of short circuiting inside the column as the cross section increased. The first 36 inch square unit was tested in parallel with the proven 18 inch column. If the underflow of the 36 incher was not as good as that of the 18 incher, short circuiting was a possibility in the larger unit. It had been designed for insertion of drop in partitions, four feed points had been provided and the 36 inch column would have become a modular unit of four 18 inch columns. Testing showed the underflows of both columns to be identical. There was no short circuiting in the 36 inch column. Once we had the 36 incher, we had no fear of the 72 inch column. It is permanently partitioned into four 36 inch columns but uses only one set of instrumentation. All the component parts of the 72 inch column come from the 36 inch unit. In spite of our results, the mining community did not believe the column could work. Finally, in 1980, Mines Gasp6 ordered an 18 and 36 inch column for their byproduct molybdenum

Jan 1, 1988

By Bee R. Waples

The Alimak Raise Method, using the Alimak Climber, has proven to be a new, fast, economical method of driving raises at the Iron King Branch of Shat-tuck Denn Mining Corporation, Humboldt, Arizona. In order to fully understand the advantages of the Alimak Raise Climber, it is necessary first to describe the basic machine; second, a complete cycle of a raise round; and third, the comparative speed and economy afforded the Iron King operation by using this new method of raise driving: The raise climber consists of a drilling platform that is raised and lowered by two 6.2 HP compressed air motors that drive, through roller chains and sprockets, three worm gears connected in series. These worm gears drive pinions which grip the guide rail rack. The worm gems are fitted with spring-loaded couplings so that a small rock or grit in the rack is passed over without disturbing the mesh between the rack and the other pinions.

Jan 1, 1961

By Gill G. Geesey, Michael R. Hankins, Paul L. Wichlacz, Anita Gianotto, John G. Jolley, Randy B. Wright

Thin films of copper [2.0 nm on germanium internal reflection elements (IREs) and 3.4 nm on germanium discs] were exposed to 10% gum arabic (aqueous solution) , 2% alginic acid (aqueous solution), 1% bacterial culture supernatant (BCS, simulated seawater solution) and 0.5% Pseudomonas atlantica exopolymer (simulated seawater solution). The IRES were monitored in situ, in real time using fourier transform infrared/cylindrical internal reflection/attenuated total reflection spectroscopy as a function of time at ambient conditions. The discs were characterized (pre- and post-exposure) by X-ray photoelectron and Auger electron spectroscopies. Ancillary graphite furnace atomic absorption spectroscopy was used to monitor the removal process of the copper thin film from the germanium substrates. Results indicate that Cu was oxidized by gum arabic, alginic acid and BCS. Furthermore, Cu was removed from the Cu/Ge interface by all four polymers. The Cu was found associated with the polymer solutions.

Jan 1, 1989

By Andrew M. Wala

The determination of air quantity needed to dilute the exhaust gases to a safe level has to be individually evaluated for particular cases. The diesel powered machines are doing some specific operations in the cyclic order required by the technical process. The dynamics of this process in certain conditions has an essential influence on environmental health conditions in the working area. The main purpose of this paper is to show, using computer simulation, how the movement (dynamic) of diesel vehicles effects environmental health standards and also how to determine the required air quantity for certain conditions

Jan 1, 1983

By J. E. Shearer

The state of Arizona produced 253,276.2 metric tons (279,000 st) of crude gypsum during 1986. This represents approximately 1.7% of the total U.S. production. Four mines provided Arizona's gypsum, three of these mines are in the San Pedro Valley. The uses of the gypsum produced in Arizona are for manufacture of cement, agricultural uses, and wallboard manufacture. The gypsum is used as a retarder in the manufacture of Portland cement. The use in agriculture is treatment of alkali soils to allow water penetration. Other deposits in the San Pedro Valley have production potential, providing there is sufficient market for additional wallboard. The author completed an evaluation of one property earlier this year.

Jan 1, 1988

By H. W. Hintze

Coal handling facilities, such as utility plants, coal terminals, ports, manufacturing, processing, etc. Protect themselves against interruptions in coal receipts by maintaining a 60 to 90 day supply in storage. Not too long ago, these storage piles were referred in terms of hundreds or thousands of tons. Today, it is not unusual to have hundreds of thousands or millions of tons in storage. Burn, or shipping and receiving, rates have increased dramatically and storage piles have grown in horizontal as well as vertical dimensions. Fixed systems for delivering and retrieving coal from storage have increased in length and operating capacity and have become highly sophisticated in their operation. With few exceptions, mobile material handling equipment is used to handle all coal beyond the reach of these systems. These machines take excess loose coal from the fixed system to the storage area, place it on the storage pile, compact it, maintain the pile, protect against spontaneous combustion, the spread of fires if they occur, and return the coal to the reclaim system as required, plus many of the small chores around the facility.

Jan 1, 1981

By Donald G. Foot

The Bureau of Mines, U.S. Department of the Interior, devised a direct flotation method for treating low-grade carnallice (KC1.MgC12. 6H20) and sylvinite (a mixture of KC1 and NaC1) ores characterized by a high content of water-insoluble impurities. The procedure consists of (1) flocculation-depression of the insoluble slimes, (2) decomposition leach (carnallite only), (3) potash rougher, cleaner and recleaner flotation, and (4) final product leaching. The technique eliminates the necessity of removing the water-insoluble slimes by mechanical or flotation desliming. Laboratory results demonstrated that the direct flotation procedure, when compared with conventional techniques, yielded equivalent or improved flotation results with equal or less reagent consumptions. The direct flotation method recovered 81.6% of the solid-phase KC1 from carnallite ore and 79.7% of the KC1 from sylvinite ore in final products containing 60.5% and 60.9% K20, respectively. A flowsheet incorporating the described flotation procedure is presented.

Jan 1, 1981

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

Ground water inflow into a mine and the re¬sultant impact on district hydrology can be pre¬dicted on a site specific basis once the ground water flow systems at the site are sufficiently well under¬stood so that a valid conceptual model can be de¬veloped. Ground water flow systems are controlled by the hydrogeologic properties of the materials at the, site, including hydraulic continuity along flow paths, and boundary conditions. Preliminary delin¬eation of hydrostratigraphic units for conceptual models frequently can be based on surface water data; specifically, spring locations and the gain-loss reaches of streams. Quantitative predictions of mine water inflow and its impact on regional hydrology require col¬lection and analysis of numerous types of data. Hydraulic conductivity must be defined quantita¬tively both parallel and perpendicular to bedding as well as areally for each hydrostratigraphic unit. The characteristics of the Davis shale in case his¬tory number 2 illustrate this concept very well. The storage characteristics also must be determined for each hydrostratigraphic unit and the potential or hydrostatic head should be quantified vertically and areally for each unit. The potentiometric measure¬ments in the Bonneterre Formation as discussed in case history number 2 illustrate this point. The data collection and data analysis techniques constitute the weak link in the analysis of mine inflow and district impact. The technology for char¬acterizing a mine site in the field has lagged behind the technology for mathematical simulation; we at¬tribute this development to the fact that computer modeling is relatively inexpensive compared to field data collection and analysis. On-site tests provide the best means of quan¬tifying hydraulic conductivity parallel to bedding and the storage characteristics of the unit. Pump tests with multiple observation wells provide the best information from on site tests. Hydrogeologie boundaries can be located from such multiple well tests. Multiport piezometers provide a means for determining the vertical potential in a single borehole. Borehole geophysics can facilitate the def¬inition of hydrostratigraphic units. Vertical hy¬draulic conductivity is difficult but not impossible to quantify with existing field test methods; labo¬ratory tests of cores fail to give a value represent¬ative of the field situation. The storage characteristics of a confined (artesian) hydro¬stratigraphic unit can be difficult to quantify when pumping changes the confined unit to an uncon¬fined (water table) case after depressurization and dewatering. Transmissivity and storage coefficient can be approximated to some extent for given field con¬ditions by computer modeling if enough is known about hydrogeologic boundaries. Modeling can be used in a trial-and-error approach to match transmissivity and storage coefficients to existing water level data corresponding to estimated discharge rates. A solution can also be obtained using a tech¬nique referred to as the inverse method. Unfortu¬nately, such a solution is seldom unique. Fracture flow models exist but practicality usu¬ally necessitates reversion to a basic Darcy flow model. Data requirements for fracture flow models could be difficult to fulfill where fracture aperture, spacing, orientation, and length are required. Ob¬taining such data at depth prior to mining would be impossible for any mine simulation in fractured media. We are of the opinion that most fractured media simulations can be accomplished using con¬ventional Darcy flow models by viewing the field problem on a macroscopic scale. The real challenge is to delineate hydrostratigraphic units properly in the field. The question of whether a broken-up fault zone is an aquifer or a barrier boundary is seldom simple. The fractures usually can be treated as an equivalent porous medium in either case. Only a few techniques exist for predicting un¬derground mine water inflow. These techniques fre¬quently are designed to evaluate shaft or tunnel inflow under steady-state head conditions in a sin¬gle layer rather than in multiple hydrostratigraphic units. Few computer programs exist that have the

Jan 1, 1986

By R. G. Johnston

Minerals have many uses in paint systems. These roles include use as primary and extender pigments, sheen control, film durability enhancement, and rheology control. These and other uses of minerals in paints, and the effects of mineral quality variations on coatings performance are discussed. As coatings performance requirements become more demanding, tighter consistency of important mineral properties, such as particle size distribution and purity, is becoming increasingly vital. Future directions of mineral use to meet projected environmental restrictions and performance demands on coatings, including the use of advanced minerals technologies, are examined.

Jan 1, 1993

By J. K. Shurie

Calcined high quality bauxite is a staple raw material of the abrasive industry world-wide. The author discusses the factors which affect the economic value of calcined abrasive grade bauxite, and proposes an approach to the problem of estimating the relative value of available materials.

Jan 1, 1990

By G. F. Raymond

Discussion by G.F. Raymond Knudsen's study presents two curious conclusions: • The kriging of blasthole assays can systematically overstate mill head grades by as much as 21% as a result of unbiased sample variance. • No estimation method is able to reduce this bias to a small margin. The study is based on a simulation using real data and what are presumed to be actual variogram parameters from a real deposit. Although I have no doubt that the first conclusion (21% overestimation) is correct for the author's simulation, I do not believe this represents a realistic mining situation. Over the past 15 years I have done extensive comparisons between exploration drill-hole assays, blasthole assays and mill head grades on seven major open-pit mines, including some very erratic gold deposits. Commonly, nugget effects on blasthole variograms were 10% - 20% higher than on exploration variograms. And in one extreme case, the difference was 50%. Even in the extreme case, ordinary kriging on blastholes agreed well with the mill head grade over the long term. In Knudsen's simulation, the blasthole nugget effect is assumed to be 200% higher than exploration data's. He supports this by variogram plots from each. My guess is that the apparent, large difference between these variograms results from a failure to account for the proportional effect (blasthole assays are likely from a higher-grade area). A simple check would be a comparison of the variance of exploration samples nearest blastholes. As for a nearly conditionally-unbiased estimator of a large random error, the arithmetic mean of all of the data certainly qualifies, provided there is an even data spacing. As a corollary, so does simple kriging, which would include, in this case, a large weighting to the arithmetic mean. Similarly, using a large number of samples with ordinary kriging or indicator kriging would significantly reduce the bias in the case of a large nugget effect for the variogram.[ ] Reply by H. Peter Knudsen Raymond questions two conclusions in the paper. First, he wonders whether a 21 % overestimation represents a realistic mining situation. I agree that 21 % is high, but overestimation in the range of 10% to 15% is certainly common in my experience. Furthermore, several years ago I consulted on a gold mine that was experiencing a 45% overestimation due, predominantly, to poor blasthole samples: In further questioning of the 21 % value, Raymond wonders whether the nugget effect of the blastholes is really so much larger than the nugget values of the exploration data. It is consistently larger throughout the deposit. In my experience with six Nevada gold mines, the high nugget value is not unusual. In fact, for some reason, nugget values for blasthole samples are typically about 0.0005 (opt squared). I am of the opinion that this high nugget effect observed at many gold mines is predominantly due to the inherent inadequacies of the blasthole sample and subsequent sample preparation. The second conclusion Raymond questions is the inability of the estimators tested to reduce the conditional bias. In fact, the conditional bias is extreme with the polygon estimator and greatly reduced by ordinary kriging. However, it was not eliminated. Raymond suggests using simple kriging, or perhaps a larger number of samples, to reduce the conditional bias. The technique of simple kriging may be less affected by the random errors in the data, but I did not test the technique. Using a larger amount of data presupposes that too few samples were used initially. In ordinary kriging and indicator kriging, the screen effect comes into play and ensures that samples beyond the second screen are given zero weight. Hence, increasing the sample size does not change the estimates nor the conditional bias. The main point of my paper is that the random unbiased errors (a fact of life in blasthole samples) cause a conditional bias in our estimates. The mechanics by which the conditional bias is introduced are nicely explained by Springett. My paper simply shows that the bias is also present when working with linear estimators, such as ordinary kriging, and even with nonlinear estimators, such as indicator kriging.[ ]

Jan 1, 1993

By M. Misra

The dissolution of mercury during gold cyanidation is the cause of many environmental, health and processing problems. A study was undertaken to find a suitable reagent for the removal of mercury bearing complexes from the leach liquor. Removal of mercury from synthetic solutions and process water from a gold mine was investigated using sodium sulfide, sodium and potassium dimethyl dithiocarbamates as precipitating reagents. The results have shown that efficient mercury removal could be achieved by addition of potassium dimethyl dithiocarbamate (KDTC), by formation of stable mercury compounds. The resultant mercury-potassium dimethyl dithiocarbamate precipitates are stable under present plant operating conditions and do not have any deleterious effect on gold adsorption by activated carbon. Furthermore, it is also possible to stabilize the mercury in the heap during cyanidation by adding dimethyl dithiocarbomate.

Jan 1, 1998

By Mark E. Emerson

Foreign and domestic companies find the mineral potential of this country's stern public lands, its capital markets, political stability, and "open-door" policy attractive for hard mineral exploration. Traditionally, exploration in the US has been geographically based and responsibility oriented at the district level. The 1970s brought a decentralized approach to exploration as major non-metal companies began to diversify into hard minerals. The emphasis was on capital intensive metal and energy projects involving bulk-tonnages. The recent economic downturn in metals has caused many of the new participants to withdraw from the mineral scene. The surviving companies have shifted their emphasis toward medium scale precious metal projects having high operating margins and rapid return on equity. Exploration expenditures in the United States are now principally focused on gold and silver to the exclusion of practically all other commodities. In the search for precious metals companies have scaled back their exploration efforts and have began to apply new approaches to target selection. The search is being assisted by rapidly evolving conceptual models. The change in orientation has led to numerous new discoveries. The finds are essentially "rediscoveries" that c,-re previously uneconomic. The majority are low-grade surface deposits that have involved low discovery costs and short lead times for development by low-cost mining and beneficiation methods. The emphasis on gold to the exclusion of all other commodities raises questions regarding the direction of exploration in the US for the rest of the 80s.

Jan 1, 1984

By G. L. Brooke

At present, there are two known cobalt deposits in the 48 contiguous states. This paper is a brief description of one of these deposits. It is a Mississippi Valley type deposit, but somewhat atypical in that the metal of greater interest is cobalt as opposed to lead, zinc, and copper. The history and present work are outlined along with a discussion of the geology and mineralogy.

Jan 1, 1981

Introduction The art of ripping has been around since the Romans built the Appian Way, but only in the last decade has ripping emerged as a close competitor with drilling and blasting in breaking up heavy rock formations. That is not to say ripping is as effective in every case, or even applicable in all cases, but it does provide a cost efficient option in many situations-if the optimum ripping techniques are used. Originally designed only to loosen hard pan, the ripper has evolved into a powerful tool, its capability increasing as the weight and horsepower of tractors increased. The tractor-mounted ripper extended the range of ripping because of the additional weight it brought to bear on the ripper tooth. Tandem ripping added the weight and power of a second tractor, extending the rippability range even further. Of course, the use of ripping depends on the type of material to be ripped. Not all rock can be ripped. Determination of rippability-an art in itself-is often inconclusive but there are certain universal principles that can be considered in the first step toward optimizing ripping techniques. Each of the three basic rock types, igneous, sedimentary, and metamorphic, exhibits characteristics that bear on rippability. Igneous rocks are the most difficult to rip because they lack the stratification and cleavage planes essential to ripping hard rock. Sedimentary rocks are generally the most easily ripped since they are built of layers differing in material, texture, color, and thickness. Metamorphic rocks vary in rippability with degree of lamination or cleavage caused by transformation from pre-existing rock. The condition of rock also affects its rippability. Although sedimentary rocks offer the best opportunity to rip and igneous and metamorphic the least, decomposed granite and other weathered igneous and metamorphic rock often can be ripped economically. Little or no trouble is encountered with hardpan, clays, shales, or cemented gravel. Likewise, highly stratified or laminated rocks offer good possibilities for ripping. However, thicklybedded rock formations generally must be drilled and blasted. The physical characteristics that favor ripping may be summarized as follows: • Fractures, faults, and planes of weakness of any kind; • Weathering, resulting from temperature and moisture changes; • Brittleness and crystalline nature; • High degree of stratification or lamination; • Large grain size; • Moisture permeated clay, shale, and rock formations; and • Low compressive strength. Ripper Types A second fundamental to optimal ripping is choosing the right equipment for a particular job. There are three types of rippers-the hinge, the parallelogram, and the adjustable parallelogram. In the hinge ripper the linkage carrying the beam and shank pivots on a fixed point at the rear of the tractor. The resulting arc from up and down movement causes up to 30° differential in the tip-ground angle. Thus, the tooth angle changes as it enters and proceeds to ripping depth. The parallelogram ripper's linkage maintains the same tip-ground angle regardless of tooth depth. This type of ripper has advantages over the hinge-type when ripping above maximum depth, but does not provide the aggressive tooth angle necessary for hard-to-penetrate materials. The adjustable parallelogram combines the benefits of both the hinge and the parallelogram rippers. It has the additional benefit of being able to vary the tip angle to the optimum angle of penetration. It can also be adjusted while moving, for the optimum ripping angle in any material. Several other factors that are important in selecting the right ripping equipment are down pressure available at the tip of the ripper, tractor flywheel horsepower, and the tractor's gross weight. Down pressure at the tip determines whether ripper penetration can be obtained and maintained. Flywheel horsepower rating determines whether the tractor has the power to advance the tip. The tractor's gross weight determines whether the tractor will have sufficient- to use the horsepower. The best ripping procedure depends on the job's actual conditions. While conditions, vary, it is still possible to apply certain proven techniques to a typical ripping situation. Caterpillar Tractor Co.'s Handbook on Ripping says optimum ripping efficiency can be realized if-before actual ripping begins-these questions and answers are considered.

Jan 1, 1983

By R. K. Mehta, C. W. Schultz, J. B. Bates

Introduction The Mineral Resources Institute, The University of Alabama, has for the past three years been engaged in a program to develop a beneficiation system for eastern (Devonian) oil shales. One objective of that program was to evaluate advanced technologies for effecting a kerogen-mineral matter separation. Column flotation was among the advanced technologies selected for evaluation. Early in the program it was shown that column flotation was superior to conventional (mechanical) flotation and to the other advanced technologies being evaluated. The investigation then proceeded toward the further objective of defining the optimum operating conditions for column flotation. One observation made in the course of optimization testing was that introducing the feed into the froth (above the pulp-froth interface) resulted in an improved combination of concentrate grade and kerogen recovery. This observation was reported in a previous paper (Schultz and Bates, 1989). Because the practice of maintaining the pulp froth interface below the feed point is contrary to "conventional" practice, it was decided to subject the observation to a systematic series of tests. This paper describes a recent series of tests and the results that were obtained. Experimental equipment and procedure The arrangement of the column cell and auxiliary equipment for continuous flow testing is shown schematically in Fig. 1. The feed sump [O] is filled with a sufficient volume of prepared sample to permit a large number of tests to be performed (typically 12). Past experience has shown this is necessary to control sample variability and variability in the size distribution resulting from ultra fine grinding. The feed slurry is maintained at about 20% solids and is constantly recirculated and stirred. The sample is metered from the circulating pipe by a peristaltic pump [O]. The feed slurry is diluted with reagentized water [O] by a second peristaltic pump [O]. Wash water [O], also reagentized, is supplied through a third peristaltic pump [O]. While this feed system may seem unduly complex, it does permit users to independently vary either the wash water rate or the net solids content of the cell. In the tests reported here, the feed rate and net percent solids were constant at 12.5 gms/min. and 3.3%, respectively. Diluted feed enters the column through 6.35 mm-diam (0.25 in.-diam) copper tubing and is discharged upwardly at the center of the column. Tailings are discharged through flexible tubing that can be adjusted so as to control the position of the pulp-froth interface. The column is 76.2 mm-internal-diam (3 in.-internal-diam) and 1090 mm (43 in.) high. It is made from lucite tubing and is fitted with a 51-mm-diam (2-in.-diam) fritted glass air sparger having an average pore diameter of 50 µm. In performing a series of tests, the concentrate and tailing are allowed to discharge continuously. The system is allowed to equilibrate for 30 minutes after the pulp and froth reach operating levels. Concentrate and tailing samples are taken simultaneously for timed intervals (five to 15 minutes, depending on the volume of sample desired). After sampling, a change in operating conditions is made and the system is again allowed to equilibrate. The tests to determine the effect of the pulp-froth interface level were part of a larger series of tests in which the objective was to optimize the conditions for a rougher flotation stage in a two stage circuit. The sample used in this series of tests was an Alabama shale ground to d90 = 23.1 µm and d50 = 7.9 µm. The operating conditions remaining constant in this series of tests were as follows: Column height - 1600 mm (63 in.) Air sparser - 50 µm (average pore diameter) Spray water - 130 cc/min. Feed rate - 12.5 gm/min (0.4 oz per min) (dry solids) Percent solids - 3.3% Frother (Dowfroth 250) - 45 ppm The variable test conditions are tabulated in Table 1. Positions of the pulp level (pulp froth interface) and feed entry are presented as a percentage of column height (as measured from the face of the air sparser). These test conditions are presented Fig. 2. At each of these test conditions, individual tests were performed at varying air

Jan 1, 1992