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By G. Michael Lloyd

The term phosphate beneficiation often conjures up visions of processing technologies that physically separate the phosphate rock from some impurity and do not result in a change in the chemical nature of the phosphate rock. However, when we consider hydrometallurgical processing as a phos¬phate beneficiation technique we have to include all chemical treatments that convert phosphate rock into a more useful product. With most of the phosphate rock mined being used for fertilizer pro¬duction, the best known and most widely employed hydrometallurgical process is wet process phospho¬ric acid manufacture. In general, all phosphate rocks are so insoluble that there is no physical treat¬ment that will allow the phosphate content to be used by vegetation in a timely manner. However, after chemical treatment, 100% of the phosphate is readily available for vegetation nutrition. If there is a chemical beneficiation process that could be considered the standard method of treat¬ing phosphate rock it would have to be the wet pro¬cess for phosphoric acid production. In this process finely ground phosphate is reacted with sulfuric acid to give phosphoric acid and phosphogypsum. The phosphogypsum crystals are removed from the phosphoric acid solution and discarded; the phos¬phoric acid is further processed to give the finished fertilizer product. The primary reason that this pro¬cess was adopted so widely is because the phosphogypsum solid that is formed can be easily separated from the phosphoric acid, thereby removing the calcium portion of the phosphate rock and the sul¬fate portion of the sulfuric acid. Other acids have been used to treat phosphate rock but nitric acid is the only other acid used commercially. Nitric acid gives a mixture of phos¬phoric acid and nitrates that can be further pro¬cessed into N-P fertilizers, but the range of N/P grades that can be produced is limited. Hydrochlo¬ric acid has been used to react with phosphate rock, and it can be used to generate a pure phos¬phoric acid but only after solvent extraction to separate the calcium chloride from the phosphoric acid. And then there is the problem of what to do with the calcium chloride. The acid that has received the most attention as a substitute for sulfuric acid is phosphoric acid. In this process, the phosphoric acid reacts with the phosphate rock to give a mono-calcium phosphate solution. This process does not dissolve all the impurities in the phosphate rock, and these solids can be removed from the solution before sulfuric acid is added to convert the soluble calcium to insoluble phosphogypsum. Despite all the claimed advantages, this process has not found favor. A suc¬cessful application of this process to treat some of the lower grade phosphate rocks being used today would seem to have some distinct advantages. Other beneficiation schemes have been pro¬posed or practiced on specific phosphate rocks on a limited scale. These include calcination followed by water leaching to remove lime, weak acid

Jan 1, 1999

By E. S. Forbus

A process for producing reductive bleach in-situ in kaolin slurries has been developed. The bleach is formed by mixing an iron sponge powder (-325M) with slurry treated with SO2. The process has been successfully demonstrated in production, and applied to several grades. The color, brightness, and viscosity of the bleached product is equivalent to or better than that of product bleached with commercial sodium hydrosulfite . The tested products can be bleached at a cost, which is 38-54% of those associated with the use of commercial bleach.

Jan 1, 1993

By R. L. Miller

The potential for recovery of enriched uranium from spent nuclear fuel which has been consolidated into a nuclear waste form was explored in the studies reported here. The waste form of interest is iron-enriched basalt (IEB)--a silicate-based, fused-cast ceramic. IEB samples were prepared containing fuel assembly materials that simulate the core debris of Three Mile Island Unit II;1,2 the nominal composition of the consolidated waste form is given in Table 1. Ore Types The IEB samples, or man-made uranium ores (MMUO), are in two different mineralogical forms. Each MU0 was made by melting the constituent materials and holding at about 1500°C for 1 h. The MMUOs were then heat treated to control crystallization and crystal growth and then cooled to room temperature.

Jan 1, 1983

By Patricia Pride, Mike Robison, Jon Bergenthal, Wojciech Klecan

"The Lower and Middle River Des Peres (LMRDP) CSO Storage Tunnel is the largest single component of Metropolitan St. Louis Sewer District’s (MSD) CSO control system improvements planned to meet consent decree requirements for MSD’s Lemay Service Area. The 13.7-km (8.5-mile) long, deep, hard rock tunnel will have a finished diameter of 9.1 m (30 ft). Beginning at the Lemay Wastewater Treatment Plant (WWTP), the alignment follows the channelized River Des Peres to the outfall of the existing Forest Park tubes, as indicated in Fig. 1.The project includes three construction shafts, 34 intakes and a deep, cavern-style pump station located adjacent to the WWTP. The tunnel will be driven from the Defense Mapping shaft to the Macklind shaft. Another shaft is located near the midpoint at the tunnel in River Des Peres Park.The Forest Park intake, located at the upstream end of the tunnel just beyond the Macklind shaft, is a large, tangential vortex intake structure. With a peak flow rate of 170 m3/s (6,000 cfs), it will have a 6.1-m (20-ft) diameter dropshaft to transfer flow into the tunnel.The project includes another three large tangential vortex-flow intake structures with peak flow rates ranging from 28 to 44.5 m3/s (990 to 1,570 cfs). These large intakes are located at points where three large sewers (formerly creeks) enter the River Des Peres. Dropshaft diameters for these intakes range from 3 to 3.7 m (10 to 12 ft) in diameter. The 30 remaining intakes are smaller tangential vortex-style, with dropshafts in the range of 0.8 to 1.7 m (30 to 66 in.) in diameter.Each intake structure drops flow into a deaeration chamber constructed adjacent to the tunnel at tunnel depth. Intake tunnels transfer flow from the deaeration chambers to the main tunnel. Most of the intake tunnels are short. However, two are long. The Glaise Creek tunnel will extend 762 m (2,500 ft) from the main tunnel to the Alaska Park intake, while the Rock Creek tunnel will extend 823 m (2,700 ft) from the main tunnel to the Rock Creek intake. The dewatering pump station, located at the downstream end of the tunnel, will be built inside a large, horseshoe-shaped cavern with two access shafts. Excavation of the cavern and its access shafts will be performed by the tunnel contractor. Build-out of the dewatering pump station will be performed under separate contract."

Jun 1, 2019

By M. C. Williams

The solution of the dynamic population balance model (PBM) of a closed grinding network consisting of a mill, sump, and cyclone is problematic. The fact that the dynamic PBM model equation solution to the Inverse Problem for grinding systems is degenerate or underspecified is demonstrated. Two numerical solution approaches to the Inverse Problem are used. These are: 1) providing additional constraints on breakage selection functions or 2) performing the Arbiter-Bhrany (or other) normalization of the selection functions. Actual experimental anthracite batch grinding data is used to demonstrate the non-unique functionality of the batch dynamic mill selection and breakage functions for a real physical system. The Levenberg-Marquardt algorithm for systems of constrained non-linear equations is used to solve the batch dynamic PBM grinding equations to obtain the grinding selection and breakage rate functions. Different solutions to the same PBM transport equations are provided. The mill, sump and hydrocyclone were modeled as a CSTR operating at various retention times. Batch dynamic PBM data was used to provide the mill kinetic and breakage selection function data. Two different solutions were obtained depending on the numerical solution approach. The severity of the non-uniqueness problem for dynamic grinding is demonstrated. Each solution approach to a dynamic PBM with transport, while giving the same prediction for a single batch grinding time, gives different solutions or predictions for mill composition for other grinding times. This fact makes dynamic nodal analysis and control problematic. The fact that the constraint solution approach gives a solution may suggest that normalization for closed networks is not necessary. Differences in solutions to the PBM cannot be excused away by inaccuracies in the data used to model the grinding phenomenon.

Jan 1, 1995

By G. R. Addie

While the cost of transporting raw phosphate matrix from a pit to a processing plant via slurry pipelines is more efficient than it could be by rail cars, conventional trucks or belt conveyors, it still amounts to many millions of dollars, and in many areas there still exists much room for improvement. By taking advantage of pipeline test data on different slurries, it is possible to reduce costs by optimizing pipeline diameters, operating velocities and slurry concentrations. Better wearing, more efficient pump designs are now available which, if applied correctly, can lead to significant savings. The authors discuss what benefits are possible in relation to the electrical operating and wear maintenance costs and go into some of the factors associated with taking advantage of these benefits, identifying what they see as being practically possible.

Jan 1, 1998

By Richard H. T. Garnett

In 1900, placer gold was discovered on the beach of Nome, AK on the southern coastline of Alaska's Seward Peninsula (Garnett, 1991). The site is about 800 km (500 miles) from Anchorage. Nome is connected to Anchorage only by air (daily) and by sea (summer). Total onshore placer gold production from Nome's immediate vicinity was more than 155.5 t (5 million oz). Alaska Gold Co, has continued to produce gold onshore with, until recently, two bucketladder dredges, each of 255-L (9-cu ft) bucket capacity. Small-scale marine dredging for gold has been at- tempted several times this century (Garnett, 1991). Western Gold Exploration and Mining Co. Ltd. Partnership (WestGold) is the only company to have been successful in achieving a significant offshore production. The Nome bucketladder-dredging operation from 1986 to 1990 constituted a unique marine-mining experience in the United States. \Inspiration Gold Inc., WestGold's predecessor, acquired six offshore leases in 1985. The total lease area of 87 km2 (21,551 acres) extended from the coastline to the 4.8-km (3-mile) offshore limit (Fig. 1). Those leases al- ready had been partially explored. Drilling was conducted off the shorefast winter sea ice. Therefore, it was rarely more than 1.6 km (1 mile) from the shore or in water deeper than about 9 m (30 ft). In 1985, estimated reserves comprised of a 5-m (17-ft) vertical thickness of sediments below 6.3 m (20 ft) of water. This meant a mean digging depth below the mean sea level (msl) of 11.3 m (37 ft). It was thought that the offshore Nome deposit possessed a feature that is common with many terrestrial gold placers - the secret of profitable exploitation of the marine deposit lay in maximizing the dredge throughput to minimize the unit costs. The inability of the project to withstand the prohibitively high cost of a new dredge was recognized. So Inspiration Gold purchased a used unit. This was the world's largest ocean- going bucketladder dredge (Fig. 2). Binzn was an 850-L (30 cu ft) bucket capacity dredge that was lying idle in Singapore. Originally constructed in the Jurong shipyard of Singapore in 1979, Bima was designed by Mining and Transport Engineering, a subsidiary of IHC Holland. Billiton Marine used the dredge for tin production at Pulau Tujuh in Indonesia (Dieperink, 1978) before it was laid up in mid-1985. Bima underwent a nine-week refit before departing Singapore on June 4,1986, with 800-t (880-st ) of deck cargo. It was dry-towed for 32 days to Alaska. A slower, less expensive, wet-tow at the authorized 5.5 kmthr (3 knots/hour) would have delayed the dredge's arrival in Alaska untl almost the onset of winter. The 13,000-km (8,100-mile) voyage was made on the deck of Smit Lloyd's "Giant-4" submersible barge towed by its 7.8- MW (10,500-hp) tugboat. The first season's dredging revealed several major shortcomings in Bima. Further repairs and modifications were undertaken at Tacoma, WA. The dredge returned to Nome for the June start of the 1987 summer dredging season. The refitted Bima worked offshore at Nome from 1987 to 1990. It produced 3.67 t (118,078 oz) of gold at an average recovered grade of 824 mg/m3 (0.0204 oz/cu yd). In 1991, the dredge was dismantled and sold.

Jan 1, 1997

By Lee W. Saperstein

P.L. 95-87 contains explicit demands for land-use planning in its reclamation and designation sections. Both sections require recognition of local land-use plans; however, the sections are not well integrated. Accommodations to these planning requirements may be found in suggestions based on British mineral planning practice. Better links can be created among local land-use planners, the state's permitting authority and designation office, and mine designers. Mining companies should take more interest in local land-use plans and in resource inventories; states should consider devolving to local government responsibility for accepting mine permit applications and designation of unsuitability petitions.

Jan 1, 1982

By A. E. Petersen

The United States Bureau of Mines conducted research on high-intensity magnetic separation methods for upgrading tantalum concentrates from Alabama. Process flow sheets were studied to maximize tin removal and minimize tantalum loss. Using laboratory-scale equipment, concentrates were produced which contained over 65% combined Ta2O5-Nb2O5. Up to 78% of the tin was recovered as a byproduct, containing about 70% SnO2. Only 1 to 5% of the tantalum reported with the tin byproduct. The producer subsequently incorporated high-intensity magnetic separation into its process flow sheet to provide concentrates for the National Defense Stockpile.

Jan 1, 1993

By John W. Conn, Harry C. Verakis

An explosion in an underground coal mine can have a devastating effect on the main fans used to provide mechanical ventilation. The Code of Federal Regulations, Title 30, Part 75.310 (CFR, 1992), specifies that each main mine fan shall be offset by at least 15 feet (4.6 m) from the nearest side of the mine opening unless an alternative method of protecting the fan is approved in the ventilation plan. Also, each main mine fan shall be protected by one or more weak walls or explosion doors. This paper presents venting and fan configuration designs which provide a comparable degree of safety for explosion protection. Design flexibility is enhanced and a means for consistency in safety judgements is afforded.

Jan 1, 1993

By D. R. Nagaraj

There are few organic ligands that can adsorb effectively on noble metals, especially gold, in aqueous media. Ligands containing sulfur and nitrogen donors are known to be the most effective. Dithiophosphorous acids, for example, are known to adsorb on gold under certain conditions and these, along with xanthates, are used traditionally to recover gold values by flotation from ores. These compounds, however, are not necessarily selective for gold since most of the associated sulfide minerals are also floated. Monothiophosphorous acids, on the other hand, were recently discovered to float gold values selectively from sulfide-containing ores in a mildly alkaline medium; thus the dicresyl monothiophosphate is now used commercially in a gold flotation operation in the U.S.

Jan 1, 1991

By Adrian C. Dorenfeld

Concentradora "Dana Aida" (100 t/d capacity) was built in 1971-1972 to treat dumps and ores containing mixed oxide-sulfide copper-gold-silver minerals from two veins. The veins, Don Felipe and aria are about 150 m apart, but are distinctly different mineralogically. The deposit is near San Juan de Guadalupe, Durango. Mexico. All new process equipment was manufactured locally, in or near Torreon, Coahuila. Sampling, ore testing, flowsheet design, plant design, drawings purchasing, financing, etc. was planned and directed by two engineers. Construction was executed by local miners and common labor (largely agricultural). The paper presents detailed costs for equipment and construction. The subsequent breaking-in operation is discussed.

Jan 1, 1979

By Steven W. Hunt

"A paper called “Global Experience with Soft Ground and Weak Rock Tunneling under Very High Groundwater Heads” was presented at the North American Tunneling 2006 conference (Holzhäuser et.al 2006). It summarized global experience with earth pressure balance (EPB) and slurry shield tunneling on nine projects that encountered groundwater heads ranging from 6 to 11 bar. Since then, additional projects have been completed with groundwater heads ranging from 7 to 15 bar. Of these, the high pressure tunneling experience on the Lake Mead Intake No. 3 project in Nevada and on the Eurasia Tunnel (Istanbul Strait Crossing) had the greatest lengths of tunneling at face pressure over 7 bar. Experience with sustained pressurized face tunneling at groundwater heads over 7 bar and risk management observations are the focus of this paper. INTRODUCTION High groundwater head, particularly at 7 bar or more, is a major challenge when tunnelling in high permeability ground and on drives long enough to require cutter changes. It has a strong impact on design and operation of Tunnel Boring Machines (TBMs) in order to protect the main bearing from contamination, prevent excessive groundwater inflow, ensure face stability and enable access to the cutterhead and excavation chamber for maintenance. During the past 20 years, TBM manufacturers and suppliers have made tremendous improvements that allow tunneling at groundwater pressures to 17 bar or more. The most significant risk with high pressure tunneling is generally that of excavation chamber access and potentially the cutterhead for cutter changes, cutterhead repairs and TBM maintenance. Hyperbaric interventions are possible and safety variances have been approved for saturation dive interventions at pressures as high as 17 bar (Lake Mead Intake No. 3). While high pressure hyperbaric intervention equipment is available and safety approvals can be obtained at many locations, few contractors consider the cost and risk of hyperbaric interventions over 7 bar to be acceptable as the primary means of excavation chamber access. Experience with high pressure tunneling prior to 2006 was summarized in a paper entitled “Global Experience with Soft Ground and Weak Rock Tunneling under Very High Groundwater Heads” (Holzhäuser et.al 2006). That paper provided information on 9 projects: South Bay Ocean Outfall, San Diego; Channel Tunnel (French Side), France; Storebaelt Tunnel, Denmark; 4th Elbe Tunnel, Germany; Wesertunnel, Germany; Westerschelde Tunnel, Netherlands; St. Petersburg Red Line, Russia; Nara Prefecture Water Conveyance Tunnel, Japan; and Tokyo Wan Aqua Line Tunnel, Japan – see Table 1."

Jan 1, 2016

By Bernard Scheiner

A number of wells were drilled at two phosphogypsum stacks, one at P-21 owned by IMC and the other at Agrico's South Pierce Stack. Each well was sampled, and composites representing 5-ft increments were leached with synthetic rainwater to determine the extent of metal and nonmetal migration. To simulate what was occurring in the stack, composite samples were placed in a column, so that the column represented the stack in miniature form. Leaching experiments showed that both metals and nonmetals moved down the column when water was added to the top of the column. Data obtained from P-21 showed that with time the concentration of metals and nonmetals reached equilibrium values. The concentration of Ca, SO4, F, Si, and PO4 were -600 ppm,~1,500 ppm, 32 ppm, 14 ppm, and 22 ppm, respectively. Column leaching of South Pierce material, representing 0 to 55 ft of stack, showed that after 355 in. of water, the stack was near equilibrium values down to the 33-ft level. With continued addition of water, the portion of the stack at equilibrium would continue to increase and eventually reach the values observed for P-21. The test showed that the sulfate concentration of the effluent from the stack would never reach the USEPA standard value of 250 ppm. During the leaching test, it was determined that the permeability of the phosphogypsum in the columns was 0.000103 cm/sec. This number was used in the modeling. The data obtained from the leaching experiments cast some doubt on the practice of entombment now being practiced for closure of phosphogypsum stacks. The data clearly shows that with time, the amount of material leaching from the stack is low in concentration, with only the calcium and sulfate at higher levels. When a stack is closed at an existing operating plant, allowing rainwater to fall on the stack and capturing the solution in the existing pond system would eliminate the need for covering the stack, since with time the closed stack would continue to leach and at some point reach the equilibrium values of the P-21 stack. Entombment means that the leachable material in the stack will always be there and will require monitoring forever. Allowing rainwater to decontaminate the stack may be the best environmental solution to the problem.

Jan 1, 1999

By Sandeep Kumar Jeripotula, Govinda Raj Mandela, Aruna Manglapady

Dumper operators are frequently exposed to whole body vibration (WBV) in surface mines. Surface mining activities involve the amalgamation of comparatively high intensity of vibration and extended exposure durations. Efficient risk reduction mandates knowing of important phases of a task that poses a threat to health of dumper operators. So far in India very limited studies have been reported on WBV exposure with regard to surface mines. This paper evaluates WBV of dumper operators based on ISO 2631-1:1997 Standards and European Union (EU) Directive 2002 for different phases of job cycle. Six dumpers were considered for this study and the vibration measurements were made for all the four phases of the job cycle by placing the triaxial accelerometer on the operator’s seat-surface and at the seat-back, independently. The findings of the study revealed that the haulage task (loaded travel and empty travel) remains the chief contributor to vibration exposure according to seat-surface and seat-back measurements. Maximum frequency weighted root mean square (RMS) of 1.12 m/s2 and 1.09 m/s2 were reported for empty travel task for seat-surface and seat-back measurements, respectively. For seat-surface measurements based on RMS, Z axis remains as the dominant axis of vibration for all the dumpers during haulage task, whereas for seat-back measurements, the dominant axis varies between X and Y. Exposure action value (EAV) based on RMS has exceeded the threshold value of 0.5 m/s2 for all the dumpers during loaded travel and empty travel for seat-surface as well as for seat-back measurements.

By Hugh E. Cronin, Joseph J. Coluccio

The Columbia Slough Consolidation Conduit Project is a 12 ft. 0 in. inside diameter—15 ft. 4 in. excavated diameter tunnel, 8,300 lf. in length. The tunnel is divided in two parts a 4,460 lf. reach to the West and a 3,770 lf. reach to the East separated by a flow regulation structure in the middle. The project was built for the City of Portland, Oregon Bureau of Environmental Services (BES) by the Frank Coluccio Construction Company (FCCC) Seattle, Washington under a $22,600,000.00 contract starting in 1998 and completing the contract in 2000 at a total cost to the City of $25,448,000. An increase of 12.6% from the original contract. The tunnel was constructed in an alluvial sediment that ranges up to 1,500 feet thick. The sediment varies from ?ne silty sand and sandy silt (SM-ML) to coarse sand and gravel (GP-GM) with varying quantities of cobbles and boulders. Most of the tunnel was constructed above the water table although one reach of almost 1,000 lf. had water to above the tunnel crown. The contractor, FCCC, used a Lovat 15 ft. 3 in.900 hp shielded TBM with multiple doors (4) of variable opening to excavate and support the tunnel. Tunnel supports were 5 in. WF 16 lbs. per foot at 4 ft. spacing and4 in. Douglas Fir lagging. Tunnel excavation was worked in 10 hour shifts with additional equipment maintenance and repair on the off time. The original Geotechnical Design Summary Report (GDSR) established a baseline for contractual purposes of slightly over 5 boulders (between 12 in. and48 in.) per foot of tunnel. (Table 1) At the pre-bid conference, three test pits were excavated to approximately tunnel invert, one at each end of the project and one in the middle near the flow regulation structure. The test pits encountered a few boulders but did not seem to support the five boulders per foot (varied from one boulder per foot of tunnel to eight b/lf.). Subsequent to the pre-bid meeting and the test pits, BES issued any addenda which changed the baseline boulder quantity in the GDSR to 1.7 boulders per foot of tunnel (varied from0.2 to 3.5 boulders per foot) a reduction of 67% from the original baseline. It is interesting to note the two separate tunnel reaches and how they were evaluated in the GDSR. Originally the Eastern portion of the work was projected at5.08 boulders per foot of tunnel and the Western portion at 5.30 boulders per foot of tunnel for a variance of less than 5%. In the revised quantity the Eastern portion was projected at 2.01 boulders per foot of tunnel and the Western portion at 1.42 boulders

Jan 1, 2003

By Alexander N. Winchell

"The deposits of the east end of the Mesabi Range cover very large areas, and consist of the same taconite that produces the hematite of the central and western Mesabi. It is not deeply covered with drift, but outcrops boldly in large masses. It is not concentrated locally into enriched bodies separated by leaner areas, but is all a fairly uniform, hard, banded, unaltered chert, carrying about 25 to 30 per cent iron as magnetite.The Mesabi Iron Company, organized by Hayden, Stone & Co., and under the direction of D. C. Jackling, has taken over these deposits and is building a Concentrating plant at the new town of Babbitt, about fourteen miles east of Mesaba, Minn. This plant is to have a capacity of 2,000 to 3,000 tons per day, and is based on the application to this low grade material of the methods which have made the so-called porphyry copper properties so successful.In 1916, an experimental mill was built in Duluth; containing full sized machinery and capable of handling from 100 to 200 tons daily. A considerable tonnage of material was concentrated here during 1916, 1917 and 1918. The work showed that the resulting product could be varied, almost at will, between 60 per cent and 70 per cent iron and .006 per cent to .030 per' cent phosphorus, and that this could be made on a large scale well within the limits of cost set by iron trade conditions. A small special cargo was sent down the lakes, late in 1918, assaying 63.27 per cent iron (natural and dry) and .008 per cent phosphorus; but it is assumed that large scale shipments will normally be of a Bessemer' grade carrying about 63 per cent iron and .020 per cent to .030 per cent phosphorus. This material is a sinter, carrying only traces of sulphur and titanium, with silica between 6 per cent and 10 , per cent, and alumina, lime, magnesia, and manganese, each below 1 per cent. Some of the iron in this sinter is hematite and some is magnetite, the percentages varying according to the treatment given on the sintering machine. Like the analyses, this hematite percentage is within reasonable control. Sinter made from finely ground magnetite has much the structure of coke, being firm yet porous, and gives good account of itself in the blast furnace."

Jan 1, 1920

By Yang Ge, Yuzhu Pan, Mansheng Chu, Jue Tang, Cong Feng, Lihua Gao, Zhenggen Liu

Increasing attention is being paid to laterite nickel ore as an alternative to nickel sulfide ore to act as a feedstock for nickel metal production. While laterite nickel ore accounts for a lion’s share of global nickel resources at approximately 72 percent, it accounts for only 42 percent of the world’s nickel production. In addition, the processing of nickel sulfide ore is confronted with enormous challenges due to deeper drilling requirements, higher production costs and depleting reserves. The nickel content in the nickel laterite ore associated with hydrated iron oxides and silicate minerals limits the application of physical beneficiation methods for the removal of iron. Thus, magnetic reduction roasting followed by magnetic separation has been regarded as one of the attractive routes to concentrate and separate the valuable metals from the impurities. 

By Joseph L. Walker

INTRODUCTION In looking over the diversity of the papers that are being presented during this meeting on The Industrial Practice of Fine Coal Processing one cannot help but be impressed with the progress the industry has made in beneficiating fine raw coals, and in the number of innovations that are under study to increase recovery and to improve the quality of the finest size particles produced at a preparation plant. Those who are not in our industry, who look at the technical program, might be confused, when they see that the papers include such topics as cleaning to zero and heavy media cyclone cleaning 28M x 100M; application and performance of fine coal jigs, tables, hydrocyclones, spirals, and flotation; filtration, dewatering and thermal drying; and ancilliary aspects such as fines handling, instrumentation and process control plus screening and liberation. This is an observation and not a criticism of the diversity of papers being presented. The program fits within the parameters of fine coal preparation as the term "coal preparation" is defined, but perhaps not as the term "fine coal" is defined. Admittedly these definitions are sometimes loose. However, each should be made more definitive if a meaningful fine coal preparation overview is to be possible. COAL PREPARATION DEFINED In the opinion of the writer, the term "coal preparation" should be defined as including all work activities that are performed on all components of a run-of-mine coal from the time it leaves a surface reserve or an underground portal, until all components are either ready for shipment or are stowed on, or in the earth in an environmentally acceptable manner. This definition excludes the work that is performed to improve coal quality or handleability inside the mine or pit. This work is commonly called "face" or "pit" preparation. FINE COAL DEFINED The term "fine coal" means different things to different people. In the past, anthracite finer than No. 3 Buckwheat (3/16 in.) was considered fines which often were discarded from the breaker as an unsalable by product; and bituminous coals finer that about 1/4 in. (slack) were either discarded or sold at cost to prevent an accumulation of "no bills" which referred to a product having no immediate sale. Although a standard definition of "fine coal" is yet to be accepted by the industry, a designation of coals by particle-size range has evolved through the wide use of three-circuit flowsheets, wherein the optimum size range by washer type determines the size fractions: a coarse-coal circuit for the plus 3/8 in., a -11- coal circuit for the 3/8 in. x 28 Mesh, and a fine-coal circuit for the remaining 28 Mesh x 0. Logically, the term "fine coal" should refer to some resultant size fraction or some top size by zero. This definition should include a 28 Mesh x 0 feed to a flotation cell but should exclude a 28 x 100 Mesh size fraction fed to a heavy media cyclone. Failure to include the resultant sizes in the definition precludes accounting of the full range of particle sizes present in the feed to the plant. Since the 28 Mesh x 0 fraction is treated in the final circuit of the three discreet washing circuits, now almost always used in U.S. commercial applications, and since it contains all of its resultant sizes, in the writer's opinion, it should be designated "fine coal". FINE COAL CLEANING Cleaning coal fines, which in the past was at best marginally profitable, is now an essential part of coal preparation strategy for a number of reasons : 1. Raw coals now cost more to produce and

Jan 1, 1988

By Douglas D. Donald

The financial analyst's job is to make an objective appraisal of the investment opportunities which exist within the various industries under his supervision within the context of the economic model provided by his firm and the a climate of investor sentiments. Industry analysis is critical. When a major investment trend develops, usually the stock of all companies within the specific industry react in the same direction. I should emphasize here, however, that an investment trend is influenced by the economic outlook, but economics is not the only determinant - investor psychology and arbitrary political factors may be of equal or greater importance. Part of our job is to recognize trends early and determine whether they are valid. Following industry analysis, the analyst must examine the specific companies within that industry since some companies usually react more violently than others to industry trends, whether they be favorable or adverse. To simplify the process of evaluating an investment opportunity, it is convenient to isolate three types of factors which have a bearing on every sort of business although the degree of influence of each factor may vary from industry to industry. These categories are, 1) independent or external factors insofar as they are beyond the control of corporate management, 2) interdependent factors insofar as influence may be exerted by industry-wide action, and 3) dependent factors covering the area of internal decisions which are basically under the control of the individual corporate managers. Although the title of this paper, "The Corporate Use of Cash Flow," is only one segment of the third category, it may be useful to continent on the first two categories as well. The independent factors may be either economic or governmental. The economic factors include cyclicality which may be very high in the case of the basic commodity industries or relatively low as in the case of utilities. Underlying growth rates, another independent factor, may range from the business equipment industry with an annual growth rate of 14% through the electric utility industry with a growth rate of about 7% per year to the

Jan 1, 1970