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By G. S. Scott, G. W. Jones

ECONOMIC waste caused by mine fires may become considerable,15 especially if a fire is allowed to spread or temporarily to get beyond control. It is important, therefore, to act promptly whenever a fire is suspected and to make every action count. To contribute toward its extinction effectively it is necessary to know at the earliest possible moment, and continuously thereafter, what is happening at the seat of the fire, which may be in a remote or inaccessible area, and also to recognize what may be expected as a result of any proposed combative measure. In combatting a fire the mining engineer and the chemist should cooperate, first deciding upon the course of action to be taken and then drawing upon the mining engineer's intimate knowledge of the physical structure of the fire area, the cost and time required to carry out proposed operations, and the probable efficiency with which these operations could be performed, and upon the chemist's knowledge for the results to be expected therefrom. After the initial steps to be taken have been decided, it devolves upon the mining engineer to see that they are carried out and upon the chemist to evaluate the results and to keep the mining engineer continuously informed in regard to a variety of matters-such as the composition and toxic or asphyxiating properties of the atmosphere in which the men may be working, whether the fire is increasing or decreasing in intensity and its probable temperature, the quantity of air leakage and the explosion hazards of the fire area. The purpose of this paper is to discuss what the chemist can and cannot contribute as a member of the fire-fighting personnel. The technical literature contains several hundred papers relating directly and indirectly to mine fires, and this fact alone is testimony to the complex nature of the phenomena involved in the birth, life, and final demise of a fire. A fire may be stopped (I) by preventing oxygen from reaching the coal or (2) by removing the conditions necessary for the rapid reaction of coal with oxygen; i.e., decreasing the temperature. All methods of prevention or control. necessarily must be based upon one or the other, or both, of these two principles. . The chemistry of a fire may be understood readily if the low-temperature oxidation of coal is considered as a starting point; then oxidation at higher temperatures (or combustion) ; and with these, the composition of the solid and gaseous products resulting from the oxidation or from the effects of the heat liberated. This discussion refers particularly to Pennsylvania anthracite, although much of the material applies equally well to coals of lower rank. OXIDATION OF COAL Oxidation of coal fundamentally is confined to the surface; that is, the rate at which oxygen penetrates into the interior of a solid block of coal is so slow that it may be ignored.6,16 Oxidation starts when a fresh surface is exposed, and the rate is

Jan 1, 1942

By J. A. B. Forster

W. G. WOOLF*—The paper has a wealth of data that take careful, detailed study. As has been indicated the highlights can be only touched in the paper. The design and the arrangement of the rabble teeth as well as some of the other details of operation show a very careful study and a scientific approach to the problem which in most plants using hearth roasters are matters of trial and error with empirical conclusions and, perhaps too often, just arbitrary supervisory preferences which may be erroneous. Here the problem has been approached from a scientific angle which think well merits careful study. The ability to treat the tonnages they get through these furnaces think is made easier by reason of the high sulphide sulphur in the resulting calcines which the balance of the plant—the leaching and so on—can cope with. As indicated by Mr. Weldon in his dis- cussion earlier this morning, in most zinc plants an attempt is made to have as low a sulphide sulphur content in the roaster calcine as possible. But in the operation of the Risdon plants, they are able to cope with sulphide sulphur content up to 5 pct which is far beyond the amount of sulphide sulphur that can be handled elsewhere. I think the autogenous roasting is made easier because driving out the last percentages of sulphide sulphur brings on the necessity for extraneous heat, and that is further exemplified because even at Risdon the author shows the desirability of using some oil to better the operation, particularly for uniformity of roasting results and to obtain more regularity of furnace operation. A. A. CENTER*—Regarding the hard crusts in the beds of the furnace and autogenous roasting, the paper speaks of the former being due, at least to a considerable extent, to lead sulphide in the crusts. I wonder if the author has considered the hard crusts as being due in large part to zinc sulphate. While I was developing large zinc operations we ran into hard crusts which had to be dug up. I took several lumps which were so hard they showed the marks of the rabble teeth very plainly and kept them in my office for some weeks. A visitor came along one day, and as he was having similar troubles, I told him what we were doing about ours. I started to pick up one of these lumps and to my amazement the thing crumbled in my hands whereas originally it had been so hard it wore the rabbles down very rapidly in the furnace. I sent the sample to the laboratory and found that it analyzed very high in water-soluble zinc and sulphate-sulphur. The crumbling of the lumps was evidently due to the anhydrous zinc sulphate gradually taking on moisture from the air and forming crystalline zinc sulphate seven parts water. The increase in volume would, of course, result in disintegration of the lumps. As to getting a high tonnage in the autogenous roast, that is partly because of finishing at about 5 pct sulphide sulphur. It is a real job to get down to, say, ½ pct sulphide sulphur, which is what is usually desired, and still put through a good tonnage.

Jan 1, 1950

By Steve Skelhorn

Edmonton South LRT is the latest expansion of the Subway system in Edmonton, Alberta, Canada: Consisting of twin 5.8m segmental lined tunnels, 280m long, driven by a Lovat EPB machine. The key challenges were as follows. A 6% downhill grade. Tight radius curves of 250m.Overlying structures within 3m of the tunnel. Strata consisting of sand above the water table, glacial tills and bed rock below the water table. Located within The University of Alberta Campus and adjacent to a hospital. Northern Alberta climate with winter temperatures routinely falling below—minus 30 degrees Celsius. This paper provides a case history and outlines the challenges involved in completing this project.

Jan 1, 2005

By S. Pinard, J. MacLellan, G. Ehlert

"The Kidd Creek concentrator is looking actively for alternative technology as means of reducing operating costs in their concentrate handling section. The use of pressure and vacuum filters to eliminate the multiple hearth dryers would have the most beneficial effect. The dryers which constitute the greater portion of concentrate drying costs consume about 35,000 m3 per day of natural gas.After the evaluation of various filters by Kidd's technology group, it was clear that, because of the low final moisture requirements of 5.5 to 6% H20, few filters were able to achieve this target economically. Lately attention has focused on the Andritz hyperbaric filter. Testwork has shown that target moistures can be achieved with an appreciable reduction in operating cost.1.0 INTRODUCTIONThe Kidd Creek Concentrator is located 28 kilometers east of Timmins, Ontario. The ore is hauled by the Kidd Creek Railroad to the concentrator where it is milled to produce separate copper and zinc concentrates.Projected concentrator feed for 1991 is 3,750,000 tonnes of ore grading 3.3 % Cu and 4.7 % Zn. The concentrate production is expected to be 458,080 dry tonnes of copper concentrate at a grade of 25.7 % Cu and 265,020 dry tonnes of zinc concentrate at a grade of 54 % Zn.Currently, there is active investigation toward cost reduction through implementation of new technologies. One area which has been extensively evaluated is the concentrate dewatering and drying section. In that section, the multiple hearth concentrate dryers represent the largest source of operating costs, mainly through maintenance and energy costs (electricity and natural gas). The replacement of these dryers with filtration technology that could produce the required concentrate dryness would represent substantial savings.In September of 1990 filtration testwork carried out by Andritz Ruthner showed that Kidd Creek copper and zinc concentrates could be filtered to below 6 and 8 % moisture respectively. In late May 1991 the laboratory testwork was duplicated on siteand one month later a 5 m2 pilot filter was tested on site. Based on the pilot test results, an economic evaluation was performed. It indicated that a reasonable payback could be realized if hyperbaric filters were to replace the existing concentrate drying section."

Jan 1, 1992

By R. L. Pienaar

Paper written on project work carried out in partial fulfilment of B Eng (Min. Eng.) degree A study was conducted to investigated and compare the feasibility of two single level extraction layouts for an intact part of theorebody of the De Beers diamond mine, Koffiefontein. Two layouts were chosen and the tonnage profile, development schedule and production ring blast layouts were calculated by applying the information acquired through a literature study. Current principles used on the mine were also applied. The first layout chosen is described as a simple straightforward layout with 6 production tunnels spaced 15 m apart, and is referred to as Layout A. The second layout, named Layout B, is described as a complex chevron layout with 9 production tunnels spaced 15 m apart. This layout has a rim tunnel situated in the kimberlite ground, which is called a chevron and this made the layout more complex than Layout A. By comparing the financial results of the two layouts, it was found that Layout B is a more feasible option than Layout A. Sensitivity analyses show that both layouts are extremely sensitive to the revenue of the projects. Therefore it is recommended that Layout B can prove to be a source of additional tonnage to add to the total tonnage profile of the mine and to extend the life of mine by 20 months, only if certain financial/economic factors improve. There is, however, still some areas of the project that need to be investigated.

Jan 1, 2006

By F Gervilla, J Laukkanen

The Kevitsa Ni-Cu-PGE deposit in northern Finland contains disseminated Ni-Cu-Fe sulfides, and platinum-group minerals (PGM). The 2.05 Ga-old ultramafic-mafic, funnel-shaped intrusion is intruded into a volcanosedimentary sequence. The major ore minerals are pentlandite, hexagonal pyrrhotite and troilite, chalcopyrite, cubanite and magnetite. Pyrrhotite and pentlandite are replaced by pyrite, millerite and heazelwoodite near the surface. Nickeline, maucherite and gersdorffite are locally abundant. A pilot feed sample of 575 t of the main orebody at Kevitsa had average grades 0.45 per cent Cu, 0.36 per cent Ni, 1.85 per cent S, 0.33 g/t Pt, 0.34 g/t Pd and 0.19 g/t Au, with 0.06 per cent Ni contained in silicates. Higher PGE contents (2 - 27 ppm) occur in a vertical N-S bearing fault zone. The PGM (n = 6022) mapped in specimens of four drill holes, consist of sperrylite 34.7 vol per cent, braggite and cooperite 18.5 vol per cent, geversite 17.2 vol per cent, moncheite 14.7 vol per cent, merenskyite 6.7 vol per cent, stibiopalladinite 4.3 vol per cent, sobolevskite 1.9 vol per cent, kotulskite 0.7 vol per cent, michenerite 0.6 vol per cent, keithconnite 0.3 vol per cent, rustenburgite 0.1 vol per cent, melonite, froodite, laurite together 0.1 vol per cent and only a few grains of sudburyite, irarsite, vysotskite, erlichmanite, hongishiite, Pt-Cu-alloys, three undefined Pd and Pt bismuthinides and antimonides and a Pd-Cu-bearing Pt-Ni-Fe sulfide. Most of the PGM, 62 per cent, are enclosed in silicates, 29 per cent are intergrown with sulfides, nine per cent are within sulfides. A portion of the PGE occurs in solid solution in sulfides and arsenides. Ni flotation yielded smelter-grade concentrates averaging 12.2 per cent nickel. The average sulfide Ni recovery is 78.8 per cent and total Ni recovery is 65.3 per cent. The average Pt is 10 g/t at 55 per cent recovery, Pd 7.7 g/t at 38 per cent recovery and Au 1.6 g/t at ten per cent recovery. Cu concentrates averaged 28.6 per cent Cu with an average recovery of 78 per cent and further 11 per cent in the nickel concentrate. The copper concentrate contains 4 g/t Au at 24 per cent recovery. The average Fe to MgO ratio 17.9 is well above the targeted minimum of 4.5. The pilot results show that the previously used TETA (triethylene tetramine) can be avoided and good quality, smelter-grade concentrates can be produced from the Kevitsa ore at economical levels of recovery by conventional flotation methods.

Jan 1, 2008

A description is given of preliminary work done to identify the problems inherent in electrochemical oxidation of metal sulphides, and particularly in the use of powder compacts as anodes in electrolysis. Lead flotation concentrate pressed into a compact can serve as an anode for direct electrowinning of lead according to the reaction PbS -7 Pb2+ +So +2e. No pYrometallurgical step is required, and sulphur is recovered in elemental form.In the experimental work, electrical conductivity of the compacts was achieved by addition of 5·5 per cent of graphite. After compaction at 10-20 tons/sq. in. the anodes were strong and remained intact throughout the electro-oxidation period. Perchlorate electrolyte was used, but fluosilicate electrolyte is also applicable. A sheet of pure lead formed the starting cathode.As anode potential increased above 1·0 V (v. SHE) PbS04 was formed in increasing amount. At 0·90 V, Pb2+ extraction efficiency was 85 per cent. Anode potential was kept low by electrolysing at 60°C and 15 mA/sq. cm., and by improving the porosity of the compact. Other metal sulphides (Zn, Fe, Cu) in the flotation concentrate also oxidize and release metal ions to the electrolyte, which would need to be purified prior to deposition of lead.Power requirement in a laboratory cell was about 0·24 kWh per lb of lead extracted.

Jan 1, 1972

By T. L. Joseph

Blast furnaces are most efficient thermally when the CO2 in the top gas is highest. Oxygen introduced in the air blast is converted to CO in the combustion zones. The extent to which CO, generated in front of the tuyeres, is converted later to CO2 depends upon the ratio of ore to coke in the charge and the degree to which CO2, formed by reduction of iron oxides with CO, reverts to CO by the secondary reaction C + CO2 = 2CO. About 300 to 400 lb. of carbon' per ton of pig iron is normally involved in the solution-loss reaction or direct reduction of FeO with carbon, which produces the same over-all results; that is, lower CO2 in the top gas. Rich2 has recently reported a reduction in fuel consumption at slower rates of blowing on two furnaces in spite of higher heat losses per ton of iron. Lower fuel consumption at slow operating rates was due to more complete reduction of the ore that resulted in higher CO2 in the top gas and consequently more efficient use of carbon. Kinney3 has reported substantial savings in fuel as a result of crushing and sizing the ore charge. Improvements in practice brought about by sintering magnetite concentrates, flue dust and fine ores have been due to the establishment of conditions favoring more complete reduction of the ore in the upper part of the furnace. The extent to which oxides and fluctuations in composition in pig iron affect the quality of steel and finished product needs further study. There is evidence that more thorough preheating and reduction of the ore in the upper part of the furnace tend to give a better quality of iron. Bogatzkii4 found from laboratory work that the amount of oxygen entering electrolytic iron, melted under various slags, was directly proportional to the concentration of free FeO in the slag and inversely proportional to the basicity of the slag. Herty and Gaines5 concluded that the silicate content of open-hearth steel before deoxidation is related to the concentration of silica in the pig-iron charge, indicating that most of the persistent inclusions are silica from the iron. They also found that

Jan 1, 1936

By W. T. Lundy

THE forms in which sulphur is commonly found-native sulphur, sulphides of many metals and sulphates-are widely distributed throughout the world. The two first mentioned are the principal sources of sulphur. Substantial amounts of sulphur in some form are recovered from gases from industrial plants and from gases from smelters treating other minerals. In recent years, plants have been installed in the United States to recover sulphur in elemental form from sour gases."' The amount of sulphur so produced is small at the present time but undoubtedly will increase in the future. The recovery of sulphur from gases is not covered in this chapter, which is confined to: (1) native sulphur deposits of Texas and Louisiana; (2) those of Italy; (3) other scattered deposits of native sulphur; and (4) pyrites deposits found in many countries of the world.30 Sulphur, a nonmetallic element, occurs normally in well developed bipyramidal or tabular crystals of the orthorhombic system, as well as in stalactitic and earthy masses. The mineral has a distinctive yellow color, which may be dark and discolored by impurities; the streak is white; the luster is resinous; and the fracture is conchoidal to uneven. Sulphur is translucent to opaque; has a hardness of 1.5 to 2.5, a specific gravity of 2.05, and an atomic weight of 32.06.16 The molecular weight of the solid is expressed by the symbol SS and that of the liquid and vapor ranges from S2 to S8, depending on the temperature. Sulphur melts at a temperature between 2340 and 248°F, depending on its crystalline state, to a liquid of dark amber color, the boiling point of which is 832°F. The viscosity near the melting point is about that of water and increases to 50,000 times that of water at a temperature of 380°F, subsequently falling rapidly to a low figure at the boiling point. Crystallization changes slowly from rhombic to monoclinic above 205 °F and in cooling the transformation is reversed. The mineral is a very poor conductor of heat and electricity; it is insoluble in water and nearly all acids and is soluble in cold carbon bisulphide and carbon tetrachloride. It ignites in air at a temperature of 478°F, burning with a distinctive blue flame with evolution of sulphur dioxide.47

Jan 1, 1949

By T. L. Joseph

Blast furnaces are most efficient thermally when the CO2 in the top gas is highest. Oxygen introduced in the air blast is converted to CO in the combustion zones. The extent to which CO, generated in front of the tuyeres, is converted later to CO2 depends upon the ratio of ore to coke in the charge and the degree to which CO2, formed by reduction of iron oxides with CO, reverts to CO by the secondary reaction C + CO2 = 2CO. About 300 to 400 lb. of carbon' per ton of pig iron is normally involved in the solution-loss reaction or direct reduction of FeO with carbon, which produces the same over-all results; that is, lower CO2 in the top gas. Rich2 has recently reported a reduction in fuel consumption at slower rates of blowing on two furnaces in spite of higher heat losses per ton of iron. Lower fuel consumption at slow operating rates was due to more complete reduction of the ore that resulted in higher CO2 in the top gas and consequently more efficient use of carbon. Kinney3 has reported substantial savings in fuel as a result of crushing and sizing the ore charge. Improvements in practice brought about by sintering magnetite concentrates, flue dust and fine ores have been due to the establishment of conditions favoring more complete reduction of the ore in the upper part of the furnace. The extent to which oxides and fluctuations in composition in pig iron affect the quality of steel and finished product needs further study. There is evidence that more thorough preheating and reduction of the ore in the upper part of the furnace tend to give a better quality of iron. Bogatzkii4 found from laboratory work that the amount of oxygen entering electrolytic iron, melted under various slags, was directly proportional to the concentration of free FeO in the slag and inversely proportional to the basicity of the slag. Herty and Gaines5 concluded that the silicate content of open-hearth steel before deoxidation is related to the concentration of silica in the pig-iron charge, indicating that most of the persistent inclusions are silica from the iron. They also found that

Jan 1, 1936

By R Dwyer, S Rea, W J. Bruckard, N McSweeney, R Holmes

"The innate ability of microorganisms to adapt to changing or inhospitable environments has often been exploited for industrial purposes and forms the basis of many current biotechnological processes. The ability of bacteria to adapt and adhere to specific mineral surfaces ensures that the energetic or nutritional requirements of the cell are met and often serves as a survival mechanism under unfavourable?growth conditions or conditions of stress. In this process, cells may rearrange or modify the macromolecules expressed on the cell surface itself or produce extracellular molecules, modifying mineral surfaces to render them more suitable for growth. The modifications that these molecules produce on mineral surfaces may result in changes to the surface charge and hydrophobicity of a mineral. As a result, these biomolecules may have some potential application as flotation collectors or modifiers.In this work, the effect of long-term exposure of Paenibacillus taiwanensis to goethite and common gangue minerals associated with Australian iron ores on the differential expression of membrane and extracellular proteins was determined. Both groups of proteins play a role in cell adherence to mineral particles. Cells were adapted to grow in the presence of each mineral over time. The proteins expressed on the cell membrane and the extracellular proteins excreted by the cell were isolated and analysed by two-dimensional polyacrylamide gel electrophoresis (2D gel PAGE). The outcomes of these studies are presented here and potential application of these products for flotation discussed.CITATION:Dwyer, R, Rea, S, McSweeney, N, Bruckard, W J and Holmes, R, 2013. Microbes and their metabolites - potential iron ore flotation reagents, in Proceedings Iron Ore 2013, pp 369-372 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Aug 12, 2013

By K. Ananthpadmanabhan

Hydrolyzable surfactants such as fatty acids and amines undergo, in addition to micellisation and precipitation, various associative interactions in aqueous solutions to form ion molecular complexes such as acid-soap, dimers etc. depending upon the solution conditions such as pH. The interaction of these collector species with the mineral surface and with other dissolved species in solution will play a significant role in determining the interfacial processes such as adsorption and flotation. Marked differences in the hydrophobicity imparted to the mineral surface depending upon the solution conditions such as Ph can be attributed to possible differences in surface activities of the above complexes. This aspect of solution chemistry of flotation systems is discussed here for two widely used collectors namely, oleic acid and dodecylamine. Equilibrium species distribution diagrams based on the available literature .data and on estimates of energy of interactions between molecules, have been obtained for oleic acid and dodecylamine solutions.

Jan 1, 1978

By Yi Luo

The efforts to mitigate longwall subsidence influences on various residential and farming structures by many coal mining companies have been successful in the recent years (Luo, et al, 2003). The successes were largely dependent on the accurate subsidence predictions, correct assessments of potential influences, appropriate design and implementation of the mitigation measures. However, such efforts for protecting large industrial structures are still very rare. In this paper, the successful assessment and mitigation of the subsidence influences associated with a very intensive dynamic longwall subsidence process on a large workshop are presented.

Jan 1, 2005

By G. M. Schwartz

THERE has been considerable controversy regarding the structure and mineral constituents produced when iron ore is sintered. This investigation was undertaken in order to establish the fundamental reactions that take place in the sintering process by determining the nature of the resulting minerals. The writer's immediate interest in the subject arose mainly from the request of E. W. Davis for information based on a microscopic study of sinter submitted by him in May, 1926. In the author's preliminary work, samples of iron-ore sinter were secured from the Minnesota School of Mines Experiment Station and from commercial plants. These were polished by the usual methods for polishing opaque ores. In order to study these surfaces satisfactorily, it is necessary to magnify them rather highly, a range from 100 to 1000 dia. being used. For the higher magnifications oil-immersion lenses and an intense source of light are necessary. After the preliminary study had shown the problems involved, a series of small samples were sintered under carefully controlled conditions and with analyzed materials. Variations in the relative amounts of ore, coal and water were made in order to study the effect of each constituent. The attempt has been to arrive at a clearer understanding of the sintering process and the reactions involved without reference to any preconceived notion as to sinter. It is clear that the study of polished surfaces has yielded results not possible with thin sections alone. The thin sections made in this work were found unsatisfactory at magnification greater than 75, whereas magnifications of 1000 were easily studied with reflected light on carefully polished surfaces. A rather complete statement of observed facts is made so that the reader may form his own conclusions. The writer's interpretations are suggested.

Jan 1, 1929

By Charles H. White

METHODS in colorimetry are based on the assumption that the intensity of the color of a definite volume of solution is directly proportional to the quantity of the color-producing substance' present. In the preparation and examination of colored solutions there are three ways in which the depth of color may be varied. These variables are; (1) the quantity of the coloring matter, (2) the quantity of the solvent, or the dilution, and (3) the thickness or depth of that portion of the solution examined. With these three variables as a basis, three general methods in colorimetry have been devised, all of which have been used in the determination of carbon in steel. In the first method, the quantity of the solution is kept the same for both the unknown steel and the standard, and the sections compared are of equal thickness, while the quantity of the standard is varied until the color obtained is of the same intensity as that produced by a definite quantity of the unknown steel. In laboratories where this method is used there is pre¬pared and kept for permanent use a series of solutions with varying amounts of the standard, representing the percentages of carbon from the lowest to the highest demanded in that laboratory. These standard solutions are kept in bottles, or test-tubes of equal diameter, and are arranged in series from the faintest to the most intense color, so that it is a simple matter to find the place of the unknown steel in the series, when the solution is of a definite volume and is contained in a tube of the same diameter as the standard tubes. In the second method, equal quantities of the standard and of the unknown steel are dissolved, the standard is diluted to a definite volume, and the solution of the unknown steel is di-

Sep 1, 1906

By M. H. Kotze, V. Yahorava

A gold-selective ion exchange resin, containing tributylamine groups, was developed by Mintek for the recovery of gold from cyanide leach solutions. The Dow Chemical Company and, more recently, Purolite commercialized the production of this gold selective resin for use in the gold mining industry, and it is marketed as Dow XZ-91419.00 and Purolite A194 respectively. A novel application of this resin is the selective recovery and separation of AuCl4 - f rom p latinum group metals (PGMs) in chloride and nitrate solutions. The current paper describes potential benefits of this type of resin for selective recovery and upgrading of gold from low grade chloride or nitrate waste streams, and the possibility of gold separation from PGMs during their refining in chloride media.

Jan 1, 2014

By J. D. Audley Smith

Heine's " blue test" for copper, as described by the authorities generally, calls for a set of standard colors; and there has been some discussion concerning the relative superiority, for this purpose, of sulphate and nitrate solutions. The whole matter evidently hinges on the preservation of the standard colors in well-stoppered bottles. The apparatus described by Q. L. Heath* cannot be much improved, when very accurate readings are required. The following method was devised for the purpose of doing away with preserved sets of standard colors, by making a fresh standard for each batch of assays. The solutions are prepared in the usual way, each ammoniacal solution being filtered into its separate bottle, and then filled up to the containing-mark and thoroughly mixed. A similar color-bottle, with an " S " etched upon it, to distinguish it from the other bottles, which are numbered to correspond with the samples, is kept to run the standard. About 150 C.C. of water is put into it, and then the amount of acid (sulphuric or nitric) present in each determination, followed by 30 C.C. of ammonia (s. g. 0.90), which should make the mixture strongly ammoniacal. The liquid is now made up almost to the containing-mark (200 c.c.), say within 1 C.C. The lowest assay is selected first and placed alongside the standard, which, at the beginning, contains no copper. A copper-solution is then dropped into the standard from a burette ; and after each addition the bottle is well shaken and compared with the assay-sample. This operation is repeated until the two shades match exactly, when the burette is read and the assay-result is calculated. The assay next in order of depth of color is now taken and treated in the same way, and so on, until the batch is completed. It

Jan 1, 1901

By Charles H. White

Methods in colorimetry are based on the assumption that the intensity of the' color of a definite volume .of solution is directly proportional to the quantity of the color-producing substance present. In the preparation and examination of colored solutions there are three ways in which the depth of color may be varied. These variables are : (1) the quantity of' the coloring matter; (2) the quantity of the solvent, or the dilution ; and (3) the thickness or depth of that portion of the solution examined. With these three variables as a basis, three general methods in colorimetry have been devised, all of which have been used in the determination of carbon in steel. In the first method, the quantity of the solution is kept the same for both the unknown steel and the standard, and the sections compared are of equal thickness, while the quantity of the standard is varied until the color obtained is of the same intensity as that produced by a definite quantity of the unknown steel. In laboratories where this method is used there is prepared and kept for permanent use a series of solutions with varying amounts of the standard, representing the percentages of carbon from the lowest to the highest demanded in that laboratory. These standard solutions are kept in bottles or test-tubes of equal diameter, and are arranged in series from the faintest to the most intense color, so that it is a simple matter to find the place of the unknown steel in the series, when the solution is of a definite volume and is contained in a tube of the same diameter as the standard tubes. In the second method, equal quantities of the standard and of the unknown steel are dissolved, the standard is diluted to a definite volume, and the solution of the unknown steel is di-

Jan 1, 1907

By John E. Conley

THE slate industry of the United States has shown a marked decline in value of products made annually since the peak year 1925, although there has been moderate improvement over the lean years 1932 to 1934. In 1940, the value of all slate products sold was approximately 5 ¾ million dollars for a total output of 473,450 short tons. The quantity sold was a little less and the value a great deal lower than the annual averages for the past two decades. Enormous quantities of waste materials result from the quarrying, splitting and milling of slate. Obviously, utilization of such waste material is desirable, not only from the viewpoint of economical operation but also for the purpose of conserving natural resources. One possible use is to produce expanded or puffed lightweight aggregate for use in concretes. Such concretes are strong enough for specific uses, and the advantage of lightness could be obtained in such structures as the walls and roofs of buildings, suspension-bridge roadways and various other applications where weight is a factor. Conceivably, these uses could consume an appreciable annual tonnage. USE OF WASTE SLATES FOR MAKING AGGREGATES The lightweight-aggregate industry has been discussed elsewhere4.5 and need not be covered in detail in this paper. The materials commonly utilized are naturally occurring aggregates such as volcanic cinder, tuff and pumice; by-product materials, such as ordinary cinders and air-cooled slags; and the specially manufactured products, including Haydite, Pottsco, Gravelite, Cel-seal, Superock and Waylite. Slates, however, have not been given the consideration they merit, especially in the United States. In the British Isles, some work has been done3 on certain Welsh slates by the Building Research Station to stimulate the use of these wastes. REQUIREMENTS FOR RAW MATERIALS As pointed out by Hughes,4 the light-weight-aggregate industry is segregated in regions adjacent to large cities in order to localize the market and avoid the necessity of long freight haulage. Waste slate from the deposits of eastern Pennsylvania, northern Maryland, and Virginia would meet the basic requirement of any low-priced structural raw material. One apparent advantage is attained when slate waste, normally discarded, is utilized in the manufacture of a by-product. Mining or quarrying costs usually have been assessed against the main product. Some additional grinding might be desirable, but this operation would be relatively cheap, as most slates grind easily. It is believed that virtually all slates, when properly heated, will expand to some degree. However, types that will expand from 3 to 7 times their original volume are preferred. Fig. I illustrates the expansion of a Pennsylvania slate. The cube originally

Jan 1, 1942

By Becerra J. O

HYL developed a direct reduction process (DR) that does not require a reformer; it is called self-reforming process. Typically, direct reduction processes based on natural gas utilise a reformer unit to transform the natural gas into reducing agents (hydrogen and carbon monoxide). The reforming reactions are: CH4+H20-?3H2+CO CH4 + CO2 - 2H2 + 2CO These reforming units require a catalyst agent to promote the kinetics of the reforming reactions. The catalyst for these reactions is typically a metallic element such as nickel in a ceramic matrix; iron is also an excellent catalyst for the reforming reactions.

Jan 1, 1999