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By Scott Honan, Tassos Grammatikopoulos, Mike Johnson, Kelton Smith, Niels Verbaan, Ernesto Bourricaudy, Eric Larochelle, Rick Sixberry

"NioCorp is developing the niobium/scandium/titanium Elk Creek carbonatite resource deposit. The carbonatite phase contains significant niobium (0.71% Nb2O5), titanium (2.68% TiO2) and scandium (72 g/t) and is composed predominantly of calcite, dolomite and ankerite. Niobium is mostly (80%) contained in the pyrochlore and the balance in various Fe-Nb-Ti oxides, which also host most of the titanium. Scandium is distributed homogeneously throughout the ore mineral assemblage. The process flowsheet recovers separate niobium, titanium and scandium products. Scandium is extracted from whole ore hydrochloric acid leach solutions through solvent extraction, and is further refined through a re-leach and precipitation process. Scandium leach residues are treated with sulphuric acid to recover separate niobium and titanium precipitates. Both hydrochloric acid and sulphuric acid reagents are recycled within the flowsheet. This paper will discuss key bench and pilot test results (focusing on the scandium flowsheet) and will present a conceptual flowsheet developed by the project team for NioCorp at SGS Minerals.INTRODUCTION NioCorp is developing North America's only niobium/scandium/titanium project. The Elk Creek project is located near Elk Creek, Nebraska, and is the highest grade niobium project in North America and the largest prospective producer of scandium in the world. These elements are unique and valuable superalloy materials that are strategic and critical to many industries and national defense technologies. The project includes the Elk Creek Carbonatite (the Carbonatite) that intruded older Precambrian granitic and low to medium grade metamorphic basement rocks. Both the carbonatite and Precambrian rocks are interpreted to be unconformably overlain by approximately 200 meters (m) of Paleozoic marine sedimentary rocks of Pennsylvanian age. As a result of this thick cover, there is no surface outcrop within the Project area of the carbonatite, which was identified and targeted through magnetic surveys and confirmed through subsequent drilling. The available magnetic data indicate dominant northeast, west-northwest striking lineaments, and secondary northwest and north oriented features that mimic the position of regional faults parallel and/or perpendicular to the Nemaha Uplift."

Jan 1, 2017

By Lloyd A. Morley, Alan M. Christman

PART A. ELECTRICAL EQUIPMENT – GENERAL Paragraph 1. Electrical Equipment; Examination. All electric equipment shall be examined, tested, and properly main¬tained by a qualified person to assure safe operating conditions. The examinations and tests shall be performed at intervals not exceeding six (6) months unless otherwise specified in Parts A, B, C, D, and E. When a potentially dangerous condition is found on electric equipment, such equipment shall be removed from service until such condition is corrected. A record of such examinations shall be kept and made available to an authorized representative of the Secretary and to the authorized repre¬sentative of the miners in such mine. Each item of major equipment, in¬cluding substations, switchgear, load centers, and so forth except powered hand tools, shall be recorded separately. The record of such examination shall include all dangerous conditions found, corrective action taken, and name of examiner, and shall be signed by the person responsible for main¬taining the equipment in safe operating condition, who shall also be a qualified person. Paragraph 2. Electrical Work; Qualified Personnel. No electrical work shall be performed on electrical distribution circuits or equipment, except by a qualified person or by a person trained to perform electrical work and to maintain electrical equipment under the direct supervision of a qualified person. Disconnecting devices shall be locked out and prominently tagged by the persons who perform such work, except that in cases where locking out is not possible, such devices shall be opened and prominently tagged by such persons. Locks and tags shall be removed only by the persons who installed them or, if such persons are

Jan 1, 1976

By F. P. Haver

The Bureau of Mines is developing a leach-electrolysis procedure to produce Special High Grade (SHG) zinc from sulfide concentrate without the generation of sulfur oxide gases. Chlorine-oxygen leaching of the concentrate yields zinc chloride (ZnCI2) which, after purification, is electrolyzed in a fused-salt cell to obtain molten zinc metal and chlorine gas for recycle. This paper describes a laboratory-scale investigation made to develop a procedure for electrowinning zinc metal from ZnC12. Zinc chloride by itself is not a suitable electrolyte because it fumes badly above its melting point and is a poor conductor of electricity. Other salts therefore must be added to prevent fuming and improve the conductivity. Optimum results were obtained with a KC1-LiCl-ZnCl2 bath at 500° C, using a current density of 5 amp/in2 and an electrode spacing of 0.5 inch. Under these conditions, SHG zinc was produced at a current efficiency of 94.3 pct and an energy requirement of 1.16 kwhr/lb. Fused-salt electrowinning eliminates the need to handle and melt cathodes and increases the production of metal per unit cell volume through the use of higher current densities than are possible with aqueous electrolysis.

Jan 1, 1976

By R. J. Grace

Primarily, this report presents the results of direct efficiency tests which were conducted on the 200-hp Coal-Pak, hot water generator, No. 3 unit, at the Otterbein College heating plant in Westerville, Ohio. Also included is a comparison of fuel consumption of the old plant with that of the new plant, a comparison of efficiencies obtained by the indirect method on a 300-hp Coal-Pak, and the cost of gas vs coal to show the savings effected by the latter. Tests were conducted with ten different bituminous coals selected by Mr. S. A. Frye, business manager at Otterbein College. The coals, of ten different suppliers, from four different states and eight different seams, represented competitive products economically available to the heating plant. The tests were conducted by the usual plant personnel which consisted of the plant superintendent and two assistants. No outside supervision of any kind was given during the tests. The tests were made from Dec. 13, 1960, through Mar. 4, 1961. During this period, 300.03 tons of coal were burned in the No. 3 unit. This tonnage represented 25.5 pct of the total coal requirements for the 1960-61 season. Operation during the period was typical for the comfort heat load which the plant supplies. Of the 78.12 operating days, the boiler was on bank or holdfire 14.31 days, or 18.32 pct of the time. The length of individual bank periods varied from 2 hr to three days. Run periods varied from about 2 hr to 7.26 days. DESCRIPTION OF PLANT AND HEATING EQUIPMENT The new Otterbein College heating plant, completed in the summer of 1957, is located in a 50x50 ft boiler room on the lower level of a combination maintenance building and heating plant. The building is situated away from the main college complex on a hillside, which allows for on-grade delivery of coal to the three bunkers. The bunkers are located directly above and connected with the individual stoker hoppers. Fig. 1 shows an in-plant view of the Coal-Pak installations. The No. 3, hot water, unit is at the extreme right. The heating plant consists of three 200-hp Coal-Paks. Unit No. 1 is operated at 4 psi and carries the steam heating load of 14,000 sq ft EDR (steam) installed in two buildings. Provision has been made to convert these buildings to hot water heating at some future date. Since the unit under maximum load conditions operates only at about 50 pct steaming capacity, a heat exchanger was installed to convert steam to hot water. Thus, during light loads in mild weather the unit can be employed more efficiently to supply both steam and hot water and eliminate the need for operating either of the two hot water units. Units No. 2 and No. 3 serve a hot water heating circuit for a large group of 21 buildings of the College. These boilers deliver hot water at temperatures up to 200°F to the various buildings through distribution lines in tunnels. The present calculated maximum heat loss for the College is 15.195 million Btu per hr. The Coal-Paks have individual design capacities of 6.700 million Btu per hr. Table I lists the design date for the Coal-Paks. The No. 3 unit was chosen as the test unit, primarily because of its bunker location at the far end of the delivery ramp. Because of this location any excess coal above bunker capacity could be conveniently stored and isolated from the adjoining bunkers. Each bunker has a capacity of 30 to 35 tons of coal. The No. 2 and No. 3 hot water generators are equipped with automatic controls which maintain the

Jan 1, 1962

By Maria I. De Rosa

The Bureau of Mines performed experiments to determine concentrations of toxic gases and liquids, and smoke-particle diameters and diameter-concentrations evolved during degradation of PVC-derived brattice with and without nylon-woven fabrics, fiberglass, cotton, and jute brattices, and other mine materials. The samples were exposed in a 20-L chamber at temperatures of 1500 to 4500 C, with a chamber airflow of 10 L/min; 27.3 cm3/s of which was directed into a submicrometer-particle detector-analyzer. During the degradation of PVC-derived brattice, large concentrations of hydrogen chloride (TLV, 5 ppm; STEL, 100 ppm) evolved at a temperature as low as 2000 C. The minimum toxic load of 3.1 mg or 40 ppm in 50 L of dilution air, with O.l~g weight loss, may be reached at a mine face with 314,000 L of dilution air, at the 5th min, during which less than 1 kg of brattice undergoes combustion. Fiberglass, cotton, and jute brattices, pine wood dust, and coal dust showed small concentrations of carbon monoxide evolving during early stages of combustion. Hydrogen cyanide gas evolved from the PVC-derived, nylon-reinforced brattice; at a later stage of combustion, large concentrations pf carbon monoxide and dense smoke evolved from all PVC-derived brattices. Experiments to characterize smoke particles showed that marked differences exist in particle diameter-concentrations among the materials.

Jan 1, 1986

By Jalna R. Zatko

The Bureau of Mines has developed and constructed an apparatus that is capable of measuring concentrations of high-molecular-weight polymers in solution at 10 ppm and less. A means for determining low concentrations of polymeric solutions was devised that was independent of slight fluctuations in solvent quality in order to measure polymers in effluent prior to recycling of the recovered solution. The concentrations of polymers ore related to a measured pressure drop observed during turbulent flow of the solution. Low concentrations of polymers of ethylene oxide and acrylamide (frequently used as flocculants for clay waste systems) in aqueous and salt solutions have been successfully measured on the apparatus constructed by the Bureau.

Jan 1, 1981

By T. Badrujaman

In September 2005, after 30 years experiencing in operating an ESF using a falling water film shell cooling system, Antam entered the new era in operating an ESF, after completely modernized its ESF No. 2 from a falling water film shell cooling system to a copper cooling system. Further, in the beginning of 2007, Ferronickel plant No. 3, where its ESF also uses copper coolers, was successfully commissioned. By operating an ESF using a copper cooling technology, Antam has to face some challenges to establish the operation stability. Besides that, the chances to maximize the furnace efficiency are greatly opened. The shielded-arc mode of electrode tip is selected to operate an ESF under relatively high furnace power to produce metal having tap temperature below 1500? and slag having tap temperature below 1600?. It leads to electric current unbalance. Further, operating an ESF using a copper cooler system under high voltage as consequence of the furnace power raise increases metal and slag tap temperature. This condition damages furnace integrity fast. Besides that, relatively high metal temperature increases its pressure during metal tapping. It leads to failure in closing the metal tap hole. The main different ways in operating an ESF using a falling water film shell cooling system and a copper cooling system lie in the electrode tip mode, the water specification used for cooler, metal tap block maintenance and allowed metal & slag level in furnace. Since a copper cooler system allows an ESF to be operated by consuming nickel ore having ratio Fe / Ni up to eight and producing slag having relatively high Fe content in slag, an undesirable silicon reversion phenomenon becomes an obstacle in reaching the operation stability. A copper cooler technology lets Antam to take a chance to increase its capacity. Modernizing the ESF No.2 had increased its furnace power from 17 MWh to 25 MWh. Meanwhile the ESF No. 3 can be operated up to 42 MWh. This technology allows an ESF to be operated under relatively high furnace power above 25 MWh safely as well as prolong the campaign life of ESF brick. Further, lowering ESF power consumption, extending the limitation of ratio S/M in nickel ore consumed and adjusting Ni grade in ferronickel product can be gained by applying this technology.

Jan 1, 2014

By Katsutoshi Inoue

Pectic acid gel crosslinked with epichlorohydrin was found to have very high selectivity to lead(II) and iron(III) ions over zinc(II) ion compared with commercial weakly cation exchange resins containing the same carboxylic functional groups, suitable for removing small amount of impurity metal ions from waste zinc plating solutions. In order to lower the cost, new adsorption gel was prepared from orange juice residue containing pectic acid by crosslinking with epichlorohydrin, which was also found to be suitable for removing these impurities. Further, very cheap adsorption gels was prepared not only from orange but also from apple juice residues by a simple method through just saponificating pectin substances contained in these residues into pectic acid by calcium hydroxide followed by water washing and drying. These adsorption gels exhibited excellent adsorption behaviors for removing the impurity metal ions away from zinc.

Jan 1, 2003

By James S. Browning

The Bureau of Mines conducted laboratory batch and pilot plant flotation tests to determine the technical feasibility of recovering commercial-grade mica from a California micaceous schist. Best results in the laboratory batch tests were obtained using sodium carbonate and sodium silicate for pH control and quartz depression, and a combination of fatty acid and cationic reagents as the mica collector. In continuous pilot plant flotation tests, an alkaline fatty acid-cationic method recovered about 70 percent of the mica in a concentrate analyzing 95.6 percent mica.

Jan 1, 1965

By Shobha Kumari

Coal is the prime source of energy in India. It now contributes to about 55% of the total power generation and will continue to have its major share in the energy mix of the country in the foreseeable future. Thermal coals of India, known for their wide variations in nature and categorized as difficult to wash due to their typical mineral matter association with the organic matrix, need suitable beneficiation technique for quality upgrade for improving the overall economics, efficiency and pollution free power generation. A non-coking high-ash coal, collected from Vasundhara open cast mine of the Ib Valley Coalfield, Orissa, has been characterized along with its different fractions obtained in the process of beneficiation for distribution of minerals using FTIR, XRD and ash analysis. The raw coal was fractionated by size as well as by density. It has been observed that ash yields and mineralogical composition (found to be mainly kaolinite and quartz) have no appreciable change in different size fractions, indicating homogeneous distribution of minerals in the coal matrix. However, for the different density fractions the mineral composition and ash yield have significant variations. The quartz content of the beneficiated coal visualized with 34% ash yield, as mandated by the Ministry of Environment and Forest (MOEF) for the use of coal in power plants is reduced to half of that of the raw coal. As quartz is the most abrasive mineral in coal, reduction of its proportion obviously signifies considerable reduction of wear of power plant machinery when beneficiated coal is fed to utility boiler for generation of power. The composition of the major minerals quartz and kaolinite as determined, also favors formation of ash of refractory nature during combustion. Keywords: IB valley coal, beneficiation characteristics,

Sep 1, 2012

By Robert J. Mitchell

A general approach to mine backfill design, including laboratory testing procedures, is outlined in this paper. Relationships between certain design parameters are developed and the effect of other factors is discussed. It is concluded that the inclusion of backfill design calculations at an early stage of mine design can often produce over-all design economies. Not only are the data needed for backfill plant and tailings disposal area design, but also the pillar recovery method and over-all mine stability are dependent on the attainable fill properties.

Jan 1, 1979

By Ajoy Singh

A study was conducted to investigate the damage potential to underground multi-seam openings caused due to vibration generated by open-pit coal mine blasting. The objective was to recommend optimal explosives weight per delay to be detonated in open-pit blasts to ensure the safety and stability of underground openings. Sixteen blasts with varying charge weights and initiation sequences were conducted at Samleshawari open-pit mine and eighty four blast vibration data were recorded in the Hingir Rampur underground colliery in India. Analysis of vibration data showed that roof of the underground openings vibrated with higher amplitude of vibration compared to pillars and floors. The amplifications of vibration in the roof at the junction were upto 4.3 times the pillar vibration. It was observed that coal chips detached from roof when the peak particle velocity reached 131 mm/s or high. The vibration magnitude recorded was higher in upper seams openings than in below seam openings. Rapid attenuation of vibration was recorded in below-seam openings at similar locations. Appropriate particle-charge weight-distance equations were established to calculate the safe explosive weight per delay to be detonated in open-pit mine for safety of multi-seams workings of underground mine. The support for the galleries and junctions were further modified for the working seams to ensure greater safety.erground Coal Mine

Jan 1, 2012

By D. F. Gibbons

THIS investigation was undertaken in order to gain further information on the mechanism of plastic deformation of iron at temperatures between room temperature and 77.2°K, and also to contribute to our understanding of the low temperature brittleness phenomenon that occurs in both the tensile and impact tests. It was decided to study as pure an iron as possible and to use the tensile test in order to simplify the problem as much as possible. Two series of iron were used throughout the investigation: iron A, obtained from the National Research Corp. and supplied as gas-free, high purity iron; and iron B, obtained from Fast' and produced by melting iron under purified hydrogen. Typical analyses of samples of both irons, after fabrication, are given in Table I. The main difference in composition of these irons is in their nickel and oxygen contents. Nickel has a very large solid solubility in iron. However, oxygen has an extremely small solid solubility. An attempt was made to reduce the oxygen content still further by remelting under purified hydrogen. This was successful in reducing the oxygen content to -0.0006 pct. However, no results on tests of this iron are included, since difficulty was encountered in fabricating the ingot. Wire specimens 1 mm in diameter were used for the investigation. The wires were produced from the ingots, as supplied, by cold rolling and swaging, with intermediate vacuum anneals at 700°C. Care was taken to reduce preferred orientation in the wires to a minimum by keeping the percentage of reduction in area between anneals of 30 to 40 pct. X-ray transmission photographs showed no indication of preferred orientation. The final grain size of iron A was 20 to 24 grains per linear mm and of iron B, 16 to 18 grains per linear mm. All specimens were tested in the decarburized and denitrided condition unless otherwise stated. The procedure was to pass hydrogen, saturated with water vapor at 26oC, over the specimens at 730°C for 2 hr. The specimens were quenched in iced water and then annealed in a vacuum of - 5x10-6 mm Hg for one day at 560°C to remove hydrogen. Higher temperatures were used for the vacuum anneal but this temperature was found to be quite satisfactory in removing hydrogen. The specimens were slow-cooled after the vacuum anneal. Where it is stated that the specimens were carburized, the specimens were given an identical decarburizing treatment and then carburized by passing hydrogen, saturated with pure n-heptane vapor at 0°C, over the specimens at 730°C. The specimens were then homogenized at 730°C for 2 hr in dry hydrogen. After the carburizing treatment the specimens were given the same vacuum anneal as for the decarburized specimens. The carbon content was determined from the maximum value of the internal friction carbon peak.2 Apparatus Fig. 1 shows a diagram of the tensile apparatus, with half the cooling coil removed. The cooling coil enabled temperatures from 200" to 77.2oK to be obtained. The apparatus was designed to fit a standard Tinius Olsen Universal tensile machine. The apparatus consisted of the rigid plate A which carried two carefully machined hooks H. These hooks attached the plate firmly to the moving crosshead of the tensile machine. The lower grip B was rigidly suspended from the plate A by three stainless steel rods. The upper grip C was joined to the upper grip of the tensile machine by the stainless steel rod E. The lower part of the apparatus was surrounded by a Dewar vessel, which was made a sliding fit inside the brass collar D. The specimen was held in a clamp which fitted into the grip, forming a ball and socket joint (see inset a). This was to attain as nearly pure

Jan 1, 1954

By Amy McBrayer, Andrea Brickey

Area mining is a surface mining method that allows for efficient mining and contemporaneous reclamation of tabular or bedded deposits, e.g., coal. Mine planning methods for area mines differ from those of open pit metal mines. Area mines in the Powder River Basin (PRB) use various processes for mine planning at both the strategic and tactical stages. With the ever changing commodity markets, it is becoming increasingly important for mine operators to be able to react quickly to the changing business environment. Tools that allow for the development and rapid comparison of production schedule scenarios are key to ensuring agile mine operations. A survey of mine engineers responsible for mine design and production scheduling at coal operations provides clarity on the current mine design and production scheduling processes in the PRB, as well as a discussion on potential solutions to upcoming challenges. In this article, we describe various mine planning processes used by the coal producers in the Powder River Basin. The survey identifies a clear gap in the ability to create comprehensive schedules associated with limited tools and lack of engineering labor needed to frequently conduct the mine design and scheduling process.

Jul 8, 2023

By P. R. Perry

This publication is a compilation of tested methods for the analyses of titanium tetrachloride, titanium sponge, and various intermediate materials and products. They are presented in concise form without the detail usually found in publication of new methods because many of them are modifications of well known procedures. The Electrometallurgical Experiment Station, Boulder City, Nev., has been developing titanium production methods since 1945. Most of this development was on a modified Kroll Process but also included work on sodium reduction and electrolytic methods. During these years a large variety of samples were analyzed. Most of the analytical problems were solved using published methods or modifications of the same. At the start, there were no published procedures for analyzing titanium sponge using spectrochemical methods. A great deal of study was given to testing several of the well known techniques, and some of them, such as direct sparking of the metal, proved to be completely unsatisfactory. Solution-spark techniques were finally tried, and satisfactory procedures were developed for some of the more common impurities. These spectrochemical methods were then put into routine use for several months before their publication.2/ Some new modifications have been made, and techniques have been developed to determine other impurities found in titanium sponge and tetrachloride. All spectrochemical methods used in 1954 are included in this publication.

Jan 1, 1955

By W. L. Freyberger

In many papers concerned with fundamentals of the froth flotation process, authors have considered the contents of a flotation cell to be comprised of two immiscible phases, the froth and the pulp. I (Of course, each of these phases is in turn made up of several solid phases plus air and water.) This generalized concept of the flotation system can be extended by considering the cell contents to include two phases: an extracting phase, the air bubbles plus the collector, which removes a selected mineral component from the second phase, the ore pulp. Thus, there is an analogy between a flotation system on the one hand, and chemical extracting systems, such a solvent extraction system or a leaching system, on the other hand. Such extracting systems may be operated in cocurrent or countercurrent fashion. This paper includes first a brief discussion on the general nature of cocurrent and countercurrent operation as applied to froth flotation, and then presents the experimental results of flotation of a specular hematite ore by each method, using a tall oil fatty acid as collector. Cocurrent Flotation A conventional flotation circuit involving roughing, scavenging, and cleaning flotation is illustrated schematically in Figure 1. In such a circuit the col¬lector, or extractant, and new feed ore enter together at the head of the rougher circuit, and flow cocurrently to the scavenger circuit where more collector is added. According to this practice, sufficient collector is added to the head of the rougher (and scavenger) so that the final tailings are of low enough grade to be discarded. Since the flow of collector and ore is cocurrent, the process consists of single stage contact between the extracting phase and the ore pulp. If the scavenger concentrate is recycled to the head of rougher, the overall circuit has some of the characteristics of countercurrent flow. In such a case part of the extractant (collector) is added at the discharge end of the circuit, and may

Jan 1, 1971

By Walter Valery, Kristy Duffy, Alex Jankovic, Peter Holtham

Hatch’s Mine to Process Optimization projects integrate mining and processing strategies, minimizing operating costs and maximizing overall profit. The methodology involves ore characterization, blast audits, plant surveys, analysis of historical data, and integrated modelling and simulation of mining and processing. Blast designs are tailored according to ore type and the downstream process is optimized for the changed feed. This approach has increased throughput and minimized costs for many operations worldwide. The implementation of this methodology at an iron ore operation that faced challenges with increasing mining strip ratio and ore body mineralogy changes is presented in this paper. The larger stripping ratio increased mining costs and constrained the run-of-mine (ROM) delivery to the process plant. Process performance was affected as the ore mineralogy shifted from predominantly hematite to include greater proportions of magnetite. The project objective was to minimize overall cost per ton mined and processed and maximize profit. Tailoring the blast designs according to ore type and process optimization increased the throughput of all process lines. This provided the potential to shut down one of the five processing line, aligning the mine and plant capacity and significantly reducing the processing costs.

Oct 3, 2019

By Doris M. Hyde

The Bahama's petroleum-dominated minerals sector constricted again in 1985. The declines were mostly the result of a soft market demand for crude oil and petroleum products. In early 1984, it was a growing unprofitability that prompted Charter Oil Co. to cease using the 350,000-barrel-per-day Bahamas Oil Refining Co. facility that it jointly owns with Chevron Oil Bahamas Ltd. After midyear 1985, because of poor petroleum economics, Chevron announced it would close the refinery, leaving only a maintenance crew. In 1984, the refinery exported less than 26 million barrels of petroleum products. Chevron's throughput was reported as averaging about 8,000 barrels per day at the time of the announcement. The crude oil transshipment terminal also registered declining activity and financial losses in 1984 and 1985. Transshipments of petroleum in 1985 amounted to about 100,000 barrels per day, one-half as much as in 1982.

Jan 1, 1987

By M. G. Fleming

The professional mineral engineer is an individual of personal integrity who has undergone rigorous intellectual instruction in the scientific disciplines fundamental to his vocation; he has learnt to apply these disciplines to the discovery, mining and treatment of mineral resources and, through practical training and experience, has acquired competence and judgment to fulfil his undertakings. In the course of this development he has recognised and accepted a threefold responsibility to society, to his employer and to his profession

Dec 1, 1971

By B. G. Klugh

(New York Meeting, February, 1912). IN a paper before the Institute at Wilkes-Barre, Pa., June 1911,1 Mr. James Gayley discussed the application of this process to iron-bearing materials. The same author2 described the results of operations at the plant at Birdsboro, Pa. The purpose of the present paper is to present further discussion of some technical details involved in the operation of this plant, and of the theoretical and practical relations of this product to blast-furnace practice. Since Oct. 1, 1911, the Birdsboro plant has been in operation at the blast-furnace of the E. & G. Brooke Iron Co. The results have been highly satisfactory, although the material treated, being chiefly the flue-dust currently made by the furnace or taken from the accumulated stock-pile, was excessively high, and yet not uniform, in carbon, so that the desirable control or regularity of the amount of fuel in the mixture treated could not be maintained. Nothing could more forcibly demonstrate the flexibility of the process than the fact that, under these conditions, the sinter produced was, in all cases, of excellent quality. The Dwight & Lloyd process, as such, and its present highly-developed mechanical appliances were described by Mr. Gayley in the paper above mentioned. While the repetition of some particulars will be unavoidable in this paper, it will be confined to features bearing upon the metallurgical principles involved. According to the principle of the process, the finely-divided iron-bearing material to be sintered is intimately mixed with

May 1, 1912