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By J. A. Procarione

Introduction Ventilation engineers have always felt somewhat uneasy when, in the normal course of their work, they are forced to consider the flow in any mine entry as "wholly rough." Under such conditions in pipes, friction factor is no longer a function of Reynolds' number. This assumption serves to simplify the problem and overcome an evident lack of information concerning the variability of K-factors. During the summer of 1985, a unique opportunity to investigate the reaction of friction factor to changing flow rates presented itself. A Utah coal mine was temporarily directing a substantial airflow through a single entry. By regulating the quantity, it would be possible to obtain a wide range of Reynolds' numbers in an entry composed of two distinct coal thicknesses. This paper describes the conducted study. The theory covering basic fluid flows in pipes that led to the questioning of constant K-factor is presented first. Experimental procedures are then described, including the use of a view camera in determining entry area and perimeter. Finally, a discussion of the observed variations in friction factor versus Reynolds' number is given. Changes of 34% and 45% in K-factor were seen in this study for the two entry heights. The implications of ignoring the relationship in this case and a call for an expanded research effort are also presented. Theory Since a gas is a fluid, the general principles of fluid mechanics are equally applicable to airflow in a mine entry as they are to the flow of water in a pipe. The Darcy equation is, therefore, equally applicable to pressure drop calculations in both situations. This famous formula was originally derived for use in circular cross sections and may be modified for any other shape by the use of hydraulic radius (defined as the area divided by the "wetted" perimeter). The Darcy equation relates the physical parameters of the pipe (diameter, length, and resistance to flow) and fluid velocity to the energy needed to create the flow. A study by Nikuradse (1933) brought to light the suspected dependency of the friction factor term in the Darcy equation upon Reynolds' number. By performing a large number of laboratory experiments, he established some basic concepts upon which other researchers built. Chief among these contributors was Colebrook (1939). Using commercial-grade pipe, he established an empirical relationship between Darcy's friction factor and Reynolds' number that is still in use today. If Colebrook's system of equations is plotted on log-log paper (as was done by Moody, 1944), the curves have some very distinct features. Of importance to the subject at hand is the relative response of friction factor to Reynolds' number at a given relative roughness. Simply stated, a pipe can appear to be either smooth or rough, depending on the flow conditions. As the Reynolds' number increases, the friction factor decreases at a decreasing rate until it becomes nearly a constant value. At this point, the flow is said to be "wholly rough." It is upon this condition that many traditional ventilation solution methods depend. For calculating pressure drops in a mine entry, Atkinson's equation is much more popular than the Darcy equation. Both of these formulas describe the same phenomenon and so have similar forms and functions. The K-factor in the Atkinson equation is equivalent to the Darcy friction factor and can be related by simply changing units, equating the pressure drops, and canceling like terms. A generalized form of this relationship is presented below as equation 1. K = yf/28,800g = [q]f/28,800 (1) where K is Atkinson's friction factor in kg/m3 (lb - min2/ft4), Y is the specific volume in Nt/m3 (lb per cu ft), p is the fluid density in kg/m3 (slugs per cu ft), g is the acceleration due to gravity (9.8 m/s2 or 32.2 fpsz), and f is Darcy's friction factor, which is unitless. The objective behind Eq. 1 is to show the strong correlation between K and f. Since these two quantities differ only by a constant (q]/28,000), the assumed independence of K-factor and Reynolds' number is suspect in light of the known response of f. Unfortunately, the formulation does not provide information concerning the actual relationship, as entry geometry likely plays a major role in the response curve. Only an indirect indication that such a relationship might exist is provided. Previous work on this subject was performed by Thomas Falkie (1958). The study involved a laboratory model composed of sections of straight, steel duct and a range of Reynolds' numbers from 15,610 to 137,200. Results from this work show a variation in friction factor with Reynolds' number in much the same way as it was found to vary in pipes. Over the range of velocities examined, a decrease of approximately 38% in K-factor was observed. This previous experience provides additional incentives to investigate the phenomenon under field conditions. Past model studies done at the University of Utah by Ozukurt (1945), Biron (1944), and Greenhalgh (1942) present similar evidence of this variability. More recently, Kharkar's (1974) study of coal mine friction factors contains similar indications that K-factor is not constant in all situations. Taken together, the data

Jan 1, 1988

By E. Forssberg, T. Nordquist

Introduction In this technical note, pilot-scale tests with two types of gravity concentrators, the Knelson separator (Anon., 1983; Forssberg and Johansson, 1984) and the Mark VIIA Reichert spiral concentrator (Balderson, 1982; Forssberg and Johansson, 1984), are described. The experimental work was concentrated on the relatively untried Knelson separator. The materials tested have different concentration characteristics: one was an artificial ore with excellent liberation characteristics and with a wide difference in density between valuable and gangue minerals; the others include a spiral concentrate containing gold from the gravity circuit of Boliden, Sweden, an old scheelite tailing from Yxsjöberg, Sweden, and an auriferous ore, containing gold, that had not been crushed to full liberation. The latter three materials contained other heavy minerals, e.g. pyrite, that can interfere with concentration. Knelson separator The Knelson separator is a compact centrifugal separator with an active fluidized bed to capture heavy minerals. A centrifugal force equal to or exceeding 50 times the force of gravity acts on the particles. The strong centrifugal force traps the heavier particles in a series of rings located in the concentrator while gangue particles are flushed out (Fig. 1). The particular separator used in the tests has a 15-cm (6-in.) diam bowl and can process between ~70 and ~250 kg/h.

Jan 1, 1988

By J. M. Dean, R. L. Grayson

Introduction With the continued downward pressure on coal selling prices, mine management must make its operations as efficient as possible. The problem is the same one throughout the mineral industries, streamlining operations to maintain competitive min¬ing costs (Born, 1991). There has been a great deal of research into better using production manpower. But the story is different for operations support personnel. Traditionally, planning man¬power requirements are done annually, with modifications to the operating plan occurring generally quarterly. On the other hand, the demands on manpower are often dynamic, being responsive to changing productivity levels, mine conditions (including water influx, bad roof, poor floor) and unanticipated project requirements (responses to machine breakdowns or a new section setup requirement because of a drop in coal quality in another area). Such changes may impact the manpower structure significantly. At times they may cause serious labor relations problems concurrently. Realignments and the generation of instability in the workforce because of active job switching (bidding) are a result. Most responses by coal operators are reactive rather than proactive, a situation that could be mollified by more accu¬rate and detailed planning of predictable manpower require¬ments. A manpower planning tool is needed that takes into consideration such things as changing productivity levels and conditions as manhour requirements for different work areas are planned, rather than blindly allocated through traditional staffing modes. Mine Management Support System The Mine Management Support System (MMSS) is an integrated decision support system that was developed using Sun Microsystems hardware according to the contract be¬tween the US Bureau of Mines and West Virginia University. The software used in the development of MMSS included X¬ windows (Scheitler and Gettys, 1986) for the user interfaces, LASER (Raman, 1985; Hayhurst, 1990) for the expert sys¬tem and graphics components, and the C language for general programming requirements. Sun DOS was used to allow the use of IBM-compatible applications software on the Sun networks. Sun OS is the operating system. It can be viewed as an enhanced version of UNIX. One of the components of MMSS that was developed to address the many specific objectives outlined in Grayson et al. (1990) is the work scheduling system. MMSS accom¬plishes work scheduling by using a mathematical (goal) programming approach to minimize the amount of switching of workers from their regular jobs to perform other jobs (Grayson and Nutter, 1991). Manpower support program The main objective of MANSUPP is to plan a detailed manpower structure for a mine by incorporating a production schedule, condition-modified work standards and manage¬ment preferences to predict required manhours for various support activities, for example, belt moves. This is accom¬plished by gathering and processing planning information the same way a mine planner would. The program is menu¬driven with the selection of a particular menu item locating the user in the appropriate area within the spreadsheet to review/add information. All of the interactions between productivity, section conditions and management prefer¬ences and their effect on the amount of support work required must be considered. Information regarding the productivity of a particular section, the geometry of the section, work standards for individual activities and section conditions must all be gath¬ered for input into the program. In the initial stages of development, it was decided that the program be developed in an environment that mine planners were familiar with in order to accommodate the use of the program. Lotus 1-2-3 or a compatible (Quattro Pro) met this environmental need. Other reasons included ease of calculation due to built-in functions, powerful command language and macro capabili¬ties. Each of the major components of the program defined are summarized. Production schedule This area allows the user to input a production schedule for the entire mine over 12 periods. Space has been provided for five sections, advance or retreat, and two longwall sections. There are several calculations occurring in this area that do not apply to the determination of manpower, but serve as an aid in production planning. Section description The section description area of the spreadsheet is shown in Table 1. The first three items are calculated in the production schedule area of the worksheet and then transferred to this location. The remaining information is input by the user to be used in later calculations.

Jan 1, 1996

For an exploration area with a predetermined grid of data points, a construction is given for elementary blocks which subdivide the entire area without omission or overlap, and with the property that each data point is situated at the centre of its corresponding block. The construction has a useful optimal property, and more importantly avoids an error which has been made recently in more than one publication.

Jan 1, 1985

Tasker (1985) has detailed the accepted practices for constructing polygonal blocks of influerrce in the estimation of ore reserves. Further he suggests that users might fall into traps such as omission or overlap when determining the area of influence for each sample point.As an example of these types of errors Tasker (1985) describes under "Coverage ofthe Area" how 25 per cent of the exploration area will not be included in an equiangular triangular grid where hexagonal blocks of influence are used.The introductory suggestion of Tasker (1985) of "incorrect procedure" by Rudenno (1983) (and Armstrong (1983)) reiterated around the Figure 2a of Tasker (1985), is based on the brief description of the polygonal area of influence by Rudenno (1983). That description is not too different from that given by King, McMahon and Bujtor (1982).

Jan 1, 1985

By R. L. Grayson

Compliance standards for respirable dust, which are reduced because of high quartz concentrations, have become an industry-wide problem. Research on the physical and chemical character of quartz-classified particles in respirable coalmine dusts indicates that they mine exist in intermixed layers or may be coated by clay or coal material. The relative amount of intermixing varies with the coal seam being mined. If most of the quartz particles of a mine dust are "pure," then the dust may have a high fibrogenic risk potential because of truly free silica and the compliance standard should be lowered. If, however, the quartz is locked inside another material matrix or coated, a reduced compliance standard may not be appropriate.

Jan 1, 1992

By D. Panda, P. K. Sahoo, A. S. Rao

The waste phosphogypsum produced by Paradeep Phosphate Limited (PPL) at Paradeep (Cuttack, Orissa, India) consists primarily of calcium sulfate with small amounts of phosphoric acid, calcium fluoride and other harmful components. Presently these are dumped as waste. In this work, laboratory trials have been made to utilize the calcium sulfate present in the phosphogypsum to prepare usable products like ammonium sulfate, calcium carbonate and magnesium sulfate.

Jan 1, 1994

Reports on process control computer projects at the Kamioka Plant are summarized from the Japanese technical literature (see references). Fluidized Bed Roaster and Sulfuric Acid plant1 A small process control computer (Yokogawa CCS-200 with an OKITAC 5090P 8K central processor) was installed in the Kamioka smelter in the summer of 1964. A year-and-one-half installation project followed, which successfully brought the roaster and acid converters under computer control. The process plant consisted of two 130 RPD fluosolids roasters and two Monsan to acid units. The roasters are charged with lead-zinc concentrates with SO2 from the roaster processed to sulfuric acid. The computer system provides the following functions: 1) log data from the plant--the input is taken from 50 analog signals, 130 high-low digital alarm points, and 50 analog high- low alarm points; 2) set-point control on 18 control loops; 3) direct digital control (DDC) on eight control loops for which conventional controllers were difficult to adapt due to abnormal loop characteristics or delay times; 4) execution of a safe shut-down procedure

Jan 1, 1969

By Wittenberg J. C, Reid A. F

A method is described for the rapid and effective production of polished section mounts of ore and mineral particulates to the standards required for automated image analysis. This requires mounting without segregation of particles by mass or density, and it is also preferable to avoid having particles in contact so that they can be recognised as individuals. To achieve this, the particle samples are first split to an appropriate degree in a rotary riffle, and the sub-sample mechanically diluted and thoroughly mixed with an inert filler of the same size range as the particles. Just sufficient epoxy resin is added to make the mixture fluid. During setting, the ore particles are mechanically supported by the inert filler particles, thus maintaining the particles in a random three-dimensional distribution. A rapid two-stage grinding process, based on the successive use of 600 mesh silicon carbide and 1200 mesh (10 um) corundum on cast iron plates, is followed by three stages of mechanical diamond polishing on cloth laps. Hand polishing is entirely avoided, as is the use of papers.

Jan 1, 1984

By A. B. Cecala, G. W. Klinowski

Introduction The US Bureau of Mines has designed and evaluated total mill ventilation systems (TMVS) at two mineral-processing operations. Both systems have proven proficient at cost-effectively reducing respirable dust levels throughout the mills. An 11.8-m3/sec (25,000-cfm) system installed at a clay-processing mill provided about 10 air changes per hour (acph). This system reduced respirable dust concentrations by about 40% throughout the mill building. The second evaluation was performed at a silica sand operation. Tests were performed with 23.6 and 47.2 m3/sec (50,000 and 100,000 cfm) of ventilation to the mill building, corresponding to 17 and 34 acph. Average millwide respirable dust reductions were 36% and 64%, respectively. This work was done to determine general guidelines and design criteria for operations that may be interested in implementing TMVS. TMVS should be designed to draw clean make-up air in near the base of the mill. This air provides general purging and may clear some dust-laden areas throughout the mill structure. It is discharged at or near the top of the structure. Without the TMVS, dust can exit the building through open doors, windows, etc. and contaminate outside plant workers. This dust can also be recirculated back into the mill building causing additional contamination problems. TMVS is not a stand-alone system. It is a supplemental technique to assist other localized dust control systems (Mody et al., 1987). Testing Total mill ventilation systems were evaluated at two processing operations. For both tests, the primary goal was to determine the reduction in respirable dust throughout the structure. Test site No. 1. The first evaluation was performed at a clay-crushing and screening mill in New York. The mill's base dimensions were 39.6 x 9.8 m (130 x 32 ft), giving a volumetric capacity of 4250 m3 (150,000 cu ft). The TMVS was designed to provide 12 m3/sec (25,430 cfm) of ventilation air to the mill, representing about 10 acph. This was accomplished by three 4 m3/sec (8475 cfm) roof-powered exhaustors that were evenly spaced across the roof. Three wall louvers provided inlets for make-up air at the mill's base. Two evaluations were performed at this mill. The first occurred for two days in December when outside air temperatures ranged from -12.2° to 4.4° C (10° to 40° F). Temperatures with the wind chill factor dropped as low as -34° to -40° C (-29 to -40° F) during the early morning. The second evaluation was in April when outside air temperatures ranged between 10° to 26.7° C (50° to 80° F). These two evaluations verified the effectiveness of TMVS in lowering respirable-dust concentrations throughout the mill. Gravimetric-sampler respirable-dust concentrations in December ranged from 0.22 to 2.39 mg/m3 under normal conditions, compared to 0.13 to 1.55 mg/m3 with TMVS. In April, respirable dust concentrations ranged from 0.29 to 4.84 mg/m3 under normal conditions compared to 0.21 to 2.37 mg/m3 with the system on. TMVS also increased visibility throughout the mill building. Dust reductions for gravimetric and RAM-I instruments are listed in Table 1. Each value was determined by comparing the average concentration for the entire day of testing for normal and TMVS values. The mean/standard deviation values for gravimetric and RAM-1 results were 42.5/17.6 and 37.1/16.8, respectively.

Jan 1, 1996

By J. W. Leonard

Comments The venerable straight-line equation Y = Bx + A is again being rediscovered, but this time by economists, mining engineers and geologists. Two economists, William Sharpe and Harry Markowitz, from the University of Chicago and Stanford University, respectively, won the 1990 Noble Prize in economics by showing that the price of individual stocks can be related to the Dow Jones industrial average (DJIA) using the straight-line equation. According to Bill Chidester, Investment Analyst, Scientific Investment, Sharpe and Markowitz strongly emphasized the role of the beta (B) value while considering the alpha (A) value to be "pretty unimportant." Chidester claims that alpha is the "mystery" value that causes individual stocks to behave independently of popular stock-market indices, such as the DJIA. For many years, mining engineers and geologists have routinely called alpha the "nugget" or surprise element in geostatistical analyses in which a designated distance (x-axis) is related to the variance of the grade (y-axis). Because the influence of the independent variable (distance) over the dependent variable (grade variance) is ideal when alpha is equal to zero, any increase in alpha represents a loosening of influence or control within the relationship. This can be good news or bad news - good news if you are an economist sorting out off-trend stocks and bad news if you are a mining engineer trying to predict sulfur in a coal seam. With many mining trends, such as those involving the relationship of distance to the variance of grade, a mathematical relationship can be formed. In geostatistical analysis, the point where a mathematical trend ends is known as the sill. If the difference between the variance of grade at the sill and at alpha can be determined, then an estimate of the relative control for a given distance can be obtained by comparing the sill value to alpha. The extent to which alpha approaches the value of the sill provides an indication of independence or loss of trend. The alpha concept may find application with distance, time or fluid problems. The point of change can be either a change in trend or the trend limit. However, even without a clearly defined point of change or trend limit, increasing alpha values always infer loss of trend. Alpha analysis should be regarded as a valuable tool when sorting out many sets of mining data. The application of this user-friendly concept to reserves, material transport, fluid flow and other areas has been, and will continue to be, the basis of many important past and future developments in mining and geology.

Jan 1, 1996

By G. G. O. O. Uwadiale

Starting from a crude ore assaying 45.6% Fe, a fine-grained oolitic iron ore was concentrated by the oil agglomeration technique to 90% Fe recovery and 65% Fe assay. The ore required grinding to minus 5 µm to effect adequate liberation, and the technique yielded optimum results at pH 9. The roles of pH, collector, and bridging liquid concentrations on selectivity and yield of iron minerals in the concentrates are discussed.

Jan 1, 1991

By G. G. O. O. Uwadiale

Because mineral matter is finely distributed within the matrix, Nigerian Labia coal is nonresponsive to conventional processing techniques. However, testing has shown that Lafia coal can be upgraded by selective oil agglormation. The optimum oil concentration was found to be 10 mL of paraffin per 25 g of coal. A product containing 6% ash and 1.9% sulfur was obtained with a yield of 90%. Optimum results were achieved with a grinding; time of from 2.5 and 3 hrs, which produced a 98% -400 mesh product. However, further grinding wars found to be deleterious to the process. Also, high pH adversely affected concentrate yields. Utilizing kerosene as a bridging liquid, a product containing 4% ash and 0.33% sulfur was achieved at a natural pH of 6.3 with a yield of 85%. At pH 9 and 10, a product containing 4.5% ash was achieved with a yield of >90%.

Jan 1, 1997

By P. Campbell, E. M. Gardy, L. Hachkowski

Introduction The efficient process control of a uranium mill and its associated mining operation requires a nearly continuous knowledge of the uranium concentration in the ore. Generally, the approach is to use laboratory assays of grab samples from the mill feed belt. In some cases, elaborate and expensive systems have been used to ensure random sampling, but even with these systems, mass-balance discrepancies still exist. There is a requirement for an on-line instrument that can give a prompt, accurate analysis of a large portion of the feed stock. The authors have recently evaluated a laboratory system that achieves these goals using neutron activation and delayed neutron counting. The development of the on-line uranium ore analyzer is a consequence of previous work done at the Whiteshell Nuclear Research Establishment (WNRE) (Campbell et al., 1978 and 1981), and is based on the emission of delayed neutrons after the neutron irradiation of fissionable materials. The mechanism of delayed neutron emission has been described (Keepin, 1965), but briefly is as follows. The fission fragments resulting from irradiation are in an excited state. Certain of these delayed fission products, precursors, decay to a more stable state by the emission of a delayed neutron. The delayed neutrons can be divided into six groups with effective half-lives ranging from 0.2-55 sec. It is important to note that more than 50% of the delayed neutrons are emitted within the first 6 sec after irradiation; this has an influence on the design of the analysis instrument. Measurement Cycle The measurement cycle involves a neutron irradiation, a delay time, and a counting period. The laboratory demonstration model of the ore analyzer comprises a californium-252 neutron source, a source chamber, a pneumatic transfer facility, and a neutron irradiation and counting module (Fig. 1). The ore sample is placed in the irradiation chamber. The neutron source is transferred to the center of the chamber for a 5-sec irradiation while the helium-3 thermal neutron detectors are switched off. As the source returns to the storage chamber, the detectors are switched on and the delayed neutrons counted. The total assay cycle is 16 sec. The current instrument design calls for the ore to be cut from the main belt every 16 sec by a mechanical arm and dropped into the analyzer directly, or via another belt. The cycle starts when there is a sufficient mass of ore in the irradiation chamber to depress a switch and stops when the delayed neutron count is finished. The ore is then dropped through the bottom of the analyzer chamber and returned to the main belt. The counts from each cycle are accumulated and the data reduced to give the desired information. Key for Fig. 1 A-Helium-3 neutron detector B-Wax moderator for neutrons C-Irradiation chamber D-Funnel for receiving ore E-Gate F-Transducers G-Biological shield for neutron source H-Transfer tube for neutron source I-Neutron source J- Microprocessor K-Accumulator for hydraulics L-Control panel M-Electro-hydraulic servo valve to control movement of the source N-Electro-hydraulic proportional valve to control movement of the gate O-Hydraulic power unit Irradiation and Counting Equipment Irradiation of the ore is achieved using a 1 X 10 7 neutron*s-1 californium-252 neutron source. The source is contained in an iron capsule and pneumatically transferred between the irradiation/counting and storage chambers. The irradiation/counting chamber possesses 12, 600-mm (24-in.) long helium-3 thermal neutron detectors embedded in an annulus of wax (Fig. 1). Two sets of six detectors are each interfaced to a preamplifier, amplifier and single-channel pulseheight analyzer system, and the two systems are connected to a scaler. A special timing unit controls the transfer of the source and the on and off switching of the detector assembly, in sequence with the movement of the source. Results and Discussion In the experimental demonstration, using the 107 neutron.s-1 source, a 10% uranium ore was measured to ± 10% (2o total error) in 10 cycles and ±6% (2o) in 100 cycles. Since a larger source was not available, it can be statistically predicted that a 109 neutron-s-1 source could be used to measure a 0.1% ore to ± 10% (20) in 10 cycles. The uranium can occur as discrete nonhomogeneous particles in the ore, and the error caused by such a distribution of particles is ± 4% (2o), as determined by

Jan 1, 1984

By S. Shapiro, R. G. Ames

Background Research on the respiratory health of coal miners has focused predominately on the relationship between coal mine dust exposure and ventilatory function impairment (Lapp et al., 1972; Jacobsen, 1976), on the prevalence of coal workers' pneumoconiosis (Lainhart et al., 1969; Morgan et al., 1973), and on coal miner mortality (Ortmeyer et al., 1974; Rockette, 1977). Lesser attention has been paid to the effects of other mining exposures. For example, underground exposure to diesel exhaust emissions (Reger et al., 1980; Ames et al., 1982,1984). An omission in US coal mining research, however, has been identified by the US Office of Technology Assessment (OTA) as the failure to study the effects of inhaled polyvinyl chloride (PVC) thermal decomposition and combustion products (Seltzer, 1979). These products may be released in the mine environment by cable fires, cable explosions, smoldering cables, and conveyor belt fires. In addition to the OTA concern, recent findings of chronic respiratory conditions and other health complications resulting from exposure to PVC decomposition and combustion products in firefighters (Peters et al., 1974; NBS, 1980; NIOSH, 1982; Wallace, 1982), and in survivors of several large scale PVC-fueled fires (Dyer and Esch, 1976; Wallace, 1981, 1982), add immediacy to the concerns. PVC decomposition and combustion PVC decomposition is initiated at temperatures far below ignition (Hartstein and Forshey, 1974; Wallace, 1981). Hence, exposures to decomposition products from smoldering cables may be important health hazards in addition to cable insulation combustion products. Hydrochloric acid (HCL) appears to be the principle toxic substance from the thermal decomposition of PVC since up to 60% by weight is released as HCL (Barrow et al., 1979). HCL is released both as a gas and as an adsorbant on soot particles (Stone et al., 1973), thus allowing penetration of respiratory defenses. In addition, laboratory results indicate that phosgene (carbonyl chloride), a potentially lethal gas, may be produced in significant quantities from electrical arcing of wires covered with PVC insulation (Woolley, 1971; Brown and Birky, 1980). Health concerns related to PVC Toxic effects resulting from high level exposures to PVC pyrolysis and combustion products in animals (Cornish and Abar, 1969) and in humans (Wallace, 1981) is beyond question. Analyses have reported acute, or short-term, health effects resulting from exposure to PVC decomposition and combustion products (Peters et al., 1974; Polakoff et al., 1975). Recently, however, several studies have found that acute effects are often followed first by apparent healing, and then by delayed chronic effects. This pattern of consequential respiratory involvement among persons involved in major PVC-fueled fires has been noted in firefighters, specifically the 1975 New York Telephone Exchange fire (Dyer and Esch, 1976; Wallace, 1982), and the survivors of the 1977 Beverly Hills Supper Club fire (Wallace, 1981). An obstructive syndrome (low FVC and FEV1) was seen in 15% of workers three months following fighting a PVC-fueled fire (Colardyn et al., 1976). Long-term respiratory symptoms seen in studies of survivors of PVC-fueled fires include the bronchitis symptoms of cough and phlegm, and lung dysfunction (Wallace, 1981). Additional support for the thesis of longer-term respiratory effects resulting from PVC-fueled fires is derived from the finding that four persons who initially survived the Beverly Hills Supper Club fire died within nine months of the fire of severe respiratory impairment (Wallace, 1981).

Jan 1, 1986

By Timothy C. Siegel

A piled raft refers to a shallow foundation that is structurally connected to the piles, while composite ground refers to a soil-pile matrix where the piles are not structurally connected. The design objectives for both a piled raft and composite ground are (excluding special considerations such as expansive soil): (1) to provide a sufficient ultimate resistance and (2) to distribute the load into the soil-pile matrix so that the settlement experienced by the shallow foundation is within tolerable limits. A simplified model is proposed for a shallow foundation on composite ground where the foundation settlement is estimated as the sum of the downward movement of the piles plus the downward movement of the shallow foundation relative to the pile head. The proposed simplified model is applied using conventional geotechnical analyses for two hypothetical examples of shallow foundations undergoing compression settlement.

Jan 1, 2011

By J. W. Jang, I. Iwasaki

Introduction Grinding ball wear in wet grinding is due to a complex combination of abrasion, corrosion, impact and their interactions. The ball charge is continually falling from a height greater than half of the ball mill's diameter, and contact is made with ore particles or other balls. Abrasive wear occurs through penetration and plowing of media materials from the ball surface by hard ore particles (Gangopadhyay and Moore, 1985). Electrochemical interactions between the abraded and unabraded areas of the ball surfaces, as well as between the grinding media and the minerals, occur in grinding mill pulps (Adam et al., 1986). The wear depends on the media composition, the hardness and the corrosive/abrasive characteristics of the slurry. The marked-ball-wear test method was used to evaluate the performance of grinding media under nitrogen and oxygen atmospheres. Also, the wear mode was studied by observing the grinding media surface under a scanning electron microscope. In general, grinding media of high hardness and high chromium contents were found to have better corrosive wear resistance than those of low hardness and low chromium contents (Jang, Iwasaki and Moore, 1988). The purpose of this investigation was to determine the size distribution and the shape of wear debris in a study of the wear rates and wear behaviors of grinding balls. Experimental Materials The grinding ball alloys used in this investigation were: •martensitic stainless steel (MSS; 0.96% C and 16.2% Cr), •high-carbon, low-alloy steel (HCLA; 0.89% C and 0.53% Cr), and •mild steel (MS; 0.22% C and 0.09% Cr). The mineral used in this study was synthetic corundum, because it was on hand and contained no magnetic impurities. Debris A 203- mm (8-in.) diam porcelain mill was used for grinding the corundum. The mill charge consisted of: •1200 g of 12-grit synthetic corundum, •770 mL of 0.05 mole/L Na2SO4 (61% solids), and •125 steel balls of nominally 25-mm (1-in.) diam. To minimize the debris corrosion, nitrogen was introduced into the mill. Grinding with MSS, HCLA-steel or MS balls was conducted for 2 hr at 51 rpm. The grinding time using the MS ball was varied from 0.5 hr to 2 hr in an attempt to study the debris amount and shape as a function of time. The debris was collected with a hand magnet from the ground slurry and separated from the corundum particles using a Davis magnetic tube. After the separation, the debris particles were immediately washed with acetone and then filtered and dried. Debris size analysis Buckbee-Mears 3-in.-diam, micro-mesh sieves and an electromagnetic shaker were used for determining the debris size distribution. In sieving, isopropyl alcohol was used. The debris fraction collected from each sieve was washed with acetone, dried and then weighed on an analytical balance. [ ]

Jan 1, 1993

By G. T. Lineberry, L. Adler

Due to safety-imposed constraints on clearances for mobile bulk handling equipment, it is necessary to optimally select a unit based on speed and size. The unit cannot be both large and fast. Since output is proportional to both size (or capacity) and speed, the combination must be carefully examined to maximize output. A simple graphical procedure and a calculus-based solution are presented to accomplish this. While the procedure applies to all mobile bulk handlers, as an example of the solution technique, attention is focused on shuttle cars. Background Major variables To fully define a bulk handling problem, the conditions, equipment, and activities must be characterized. The conditions and activities, when lumped together, are termed "encumbered space," since they "act to impede motion, as with a burden" (Lineberry, 1979). The insertion of mobile bulk handling equipment into what is already a difficult set of conditions and activities compounds the state of encumbrance, as limited space is competed for by many normal and life support systems. Long (1972) recognized the effects of design criteria, including road width, on haulage vehicle speed and, therefore, on production. Velocity -clearance curves Velocity-clearance curves are either available or can be constructed in the field. A preliminary set of curves for average conditions and activities is shown in Fig. 1, where clearances are only for the side of the vehicle (Lineberry and Adler, 1984). These curves set the safe speed limits for a vehicle operating at given clearances, under given conditions and activities. When feasible, velocity-clearance curves should be constructed through field measurement and observation. Curve construction should also consider haulway roughness, distances, and operator skill, as well as loading and dumping characteristics. It should be noted that haulage capacity determination is a total systems problem, not one just dependent on velocity and clearances. Production requirements and company capabilities also influence capacity determination.

Jan 10, 1985

By G. T. Lineberry, L. Adler

Due to safety-imposed constraints on clearances for mobile bulk handling equipment, it is necessary to optimally select a unit based on speed and size. The unit cannot be both large and fast. Since output is proportional to both size (or capacity) and speed, the combination must be carefully examined to maximize output. A simple graphical procedure and a calculus-based solution are presented to accomplish this. While the procedure applies to all mobile bulk handlers, as an example of the solution technique, attention is focused on shuttle cars. Background Major variables To fully define a bulk handling problem, the conditions, equipment, and activities must be characterized. The conditions and activities, when lumped together, are termed "encumbered space," since they "act to impede motion, as with a burden" (Lineberry, 1979). The insertion of mobile bulk handling equipment into what is already a difficult set of conditions and activities compounds the state of encumbrance, as limited space is competed for by many normal and life support systems. Long (1972) recognized the effects of design criteria, including road width, on haulage vehicle speed and, therefore, on production. Velocity-clearance curves Velocity-clearance curves are either available or can be constructed in the field. A preliminary set of curves for average conditions and activities is shown in Fig. 1, where clearances are only for the side of the vehicle (Lineberry and Adler, 1984). These curves set the safe speed limits for a vehicle operating at given clearances, under given conditions and activities. When feasible, velocity-clearance curves should be constructed through field measurement and observation. Curve construction should also consider haulway roughness, distances, and operator skill, as well as loading and dumping characteristics. It should be noted that haulage capacity determination is a total systems problem, not one just dependent on velocity and clearances. Production requirements and company capabilities also influence capacity determination.

Jan 1, 1986

By G. H. Lewis, C. L. Krcil, N. S. Harding, D. E. Havekotte

Introduction Limestone injection work was completed in a 440-kW (1.5-million Btu per hour) pilot-scale combustor in preparation for a 50,000-kg/h (110,000-lb per hour) industrial steam boiler limestone injection test program. The main objectives were to: (1) identify the important variables that affect sulfur capture, (2) determine the maximum SO2 removal, and (3) determine the effect of limestone addition on furnace operability and heat transfer. A more detailed report is given by Lewis et al. (1985). The test program was divided into two phases. The objectives of Phase I were to determine the important variables that affect SO2 reduction and to investigate the effect of limestone addition on combustor operation and heat transfer surfaces. The objective of Phase II was to determine, using only the important variables identified during Phase I, the conditions that resulted in the greatest sulfur dioxide reduction. Phase I Test program Using one limestone, five variables were identified as having an effect on SO2 reduction. These are noted with their respective ranges: heat release rate [103 to 166 kW/m3 (10,000 to 16,000 Btu per hour-cu ft)], Ca/S molar ratio (0 to 3), flue gas oxygen concentration (3% to 5%), additive particle size [90% -150 µm (-100 mesh) and 90% -45 µm (-325 mesh)], and burner stoichiometry (SR = 0.75 and >1.15). A 2.3% sulfur northern West Virginia coal, pulverized to a nominal grind of 70% -75 µm (-200 mesh), was burned for the tests. The limestone used was obtained from a Maysville, KY, quarry and contained 51.75% (wt) CaO. An 11-run statistically designed program (plus three duplicates) was used to screen the five variables. The effect of limestone injection on the combustor heat transfer surfaces was determined by comparing the heat transfer rates through a slag panel and fouling probes during periods with and without limestone injection. Combustor operation was determined by noting any operating problems that occurred during the tests. Test procedure After furnace wall temperatures had stabilized at operating conditions, heat transfer surfaces (slag panel and fouling probes) and the convection section were blown clean, and a four-hour test was begun. The

Jan 12, 1986