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By W. Luyken

A NUMBER of articles have been published, notably those by R. S. Handy, R. T. Hancock and A. P. Watt in Engineering and Mining Journal, dealing with the calculations involved in ore dressing.. These publications demonstrate both, the importance of the subject and the lack of success in developing a generally recognized method for comparing enrichment effects. Considering that we may wish -to compare enrichment results on similar or dissimilar ores, that we may desire to make the comparison on the basis of either the absolute or the economical efficiency, and that the number of factors upon which the result depends is large, the question arises whether it is possible to set up a positive method for such calculations. In the following pages it is our aim to show that a sure, complete and successful way to overcome existing difficulties has been found. GRAPHICAL ANALYSIS OF ORE-DRESSING PROCESSES The following is from Truscott's Text Book of Ore Dressing (London, 1923) x = assay of ore treated (feed). y = assay value of the concentrate. z = assay value of the tailing. a = weight of feed. b = weight of concentrate. c = weight of tailing. C = fractional concentration = b/a= x-z/y-z (1) K = ratio of concentration = 1/C = a/b =y-z/x-z (2) R = fractional recovery = yb/xa = Cy/x. (3) 100R = percentage recovery. Copt = theoretically best concentration. 100 - Copt = waste in the feed.

Jan 1, 1929

By R. L. McNearny

A comparison study was conducted between an actual physical model and a numerical model based on the physical model of the block cave mining method. Results from the numerical analyses closely simulated the blocky behavior of the physical models, including bridging and interlocking of the simulated orebody during draw. The draw-down patterns and the rate of draw developed during numerical modeling also closely simulated the physical models. An unexpected result from the numerical modeling was the development of a zone of downward deflection which preceded the development of the zone of material movement.

Jan 1, 1997

By J. W. Stickney

"On July 1, 1977 Hudson Bay Mining and Smelting Company Limited announced the commencement of construction of a 3800 ton per day concentrator at Snow Lake, Manitoba. Snow Lake, a community of 2200 people, is located 130 road miles from the main metallurgical plants at Flin Flon.The design, detailed engineering and equipment selection for the project was awarded to Wright Engineers Limited of Vancouver, B.C., who would be working in close liason with Hudson Bay Mining personnel throughout the project.The feasibility of constructing a metallurgical plant in the Snow Lake area, in addition to the concentrator at Flin Flon, had been under study since 1964 with plants of varying capacities and designs being considered. Each successive study recommended a plant of increased capacity and sophistication and finally upon an in-depth study in 1976 of ore reserves, freight rates and operational benefits, the construction of a 3800 ton per day copper-lead-zinc concentrator appeared highly benificial."

Jan 1, 1981

By E. Stamboliadis

Dithiophosphates are reaction products of phosphorous pentasulfide with various organic compounds such as phenols, alcohols, mercaptans, dioalcohols, amines, and nitrides. The products with alcohols are commercially known as "aerofloats" and these are investigated in this presentation. It appears that the formation of insoluble metal dithio phosphates is responsible for their collecting properties with minerals." In this aspect they are very similar to xanthates.3 The solubility of metal dithiophosphates depends on the number of carbon atoms in the organic chain of the corresponding alcohol. An empirical mathematical model is presented here. It allows one to predict the number of carbon atoms necessary to ensure that a certain metal dithiophosphate will form at a selected pH value of the water solution. The Nature of Dithiophosphates Dithiophosphoric acid is the product of the reaction of phosphorus pentasulfide with an alcohol, as shown in the following equation: [P2S3 + 4 ROH - 2 (RO)2 PS2H + H2S (1)] where R stands for a hydrocarbon chain. The acid so produced can be neutralized by a base to produce the corresponding salt as shown in the following: [(RO), PS2H + NH- (RO)2 PS, NH,. (2)] In this aspect dithiophosphates are similar to xanthates, the only difference being the substitution of the tetravalent carbon atom in the xanthate polar group by a pentavalent phosphorus atom in the dithiophosphate group. This difference implies at least two characteristics: (1) that there are two hydrocarbon chains associated with the polar group so that an increase in the number of carbons of the single chain is expected to have a pronounced effect on the flotation properties, and (2) that the bond between the sulfur and the central atom is stronger than in the xanthate group. This probably leads to a decrease in both the polarizability of the electronic cloud of the sulfur and the interaction with cations.3 However, what is lost in the dithiophosphates by having a weak polar group can be gained by an increase in the number of carbons as will be seen later. Taking into account the size of the atoms involved in a dithiophosphate molecule, as listed in Table 1, It is possible to draw an approximate diagram of its structure as shown in [Fig. 1]. It is assumed here that the length of the single hydrocarbon chain is equal to that of the corresponding alcohol. Preparation and Purification Acid dithiophosphates can be prepared, as previously mentioned, in accordance with Eq. 1. The rate of this reaction is too slow at room temperature and can be improved by increasing the temperature to about 60° or 80°C7 or by evacuating the system. In our investigation evacuation was performed by a water aspiration pump. The dithiophosphoric acid so produced is liquid and dissolves in the alcohol. At the end of the reaction the product is filtered from the remaining P2S3; and dissolved in petroleum ether. By blowing anhydrous ammonia gas into the ether phase the ammonium salt forms as a white precipitate and can be sepa-

Jan 1, 1977

The following guidelines have been recommended by the Diamond Exploration Best Practices Committee established by the Canadian Institute of Mining, Metallurgy and Petroleum. These guidelines are applicable to public companies and their consultants reporting diamond exploration results. In respect of these guidelines, Diamond Exploration Results are results derived from the initial activities undertaken to investigate a diamond bearing prospect or deposit, to generate tonnage, grade and value estimates that cannot be classified as a Mineral Resource or Mineral Reserve. They do not address the earlier stage reporting of geophysical, geochemical or kimberlite indicator mineral results. Public reports of Diamond Exploration Results relating to mineralization not classified as Mineral Resource or Mineral Reserve must address the confidence limits of the result and contain sufficient information to allow a considered and balanced judgement of the significance of the results. The objective of these guidelines is to assist public companies and consultants in making appropriate disclosure of diamond exploration results by reporting in a manner which is: ??Uniform within the diamond exploration industry; ??Consistent with reporting from other industry sectors; ??Comprehensive; ??Unambiguous. The guidelines are not prescriptive. Materiality is the overriding principle that determines the information that should be disclosed. However, where a company or consultant deviates from the guidelines, the disclosure or technical report should state the extent to which the guidelines have not been followed, and the reasons for the deviation. The guidelines do not in any way replace jurisdictional reporting requirements (For example, the requirement for Canadian reporting issuers to comply with Canadian Securities Administrators National Instrument 43-101 by having technical disclosure endorsed by a Qualified Person as defined in that legislation).

May 1, 2003

By C. R. Kuzell

INTRODUCTION KNOWLEDGE of the thermal capacity of gases is of great importance in making metallurgical calculations. The metallurgist is, frequently called upon to investigate and determine furnace efficiencies in which the heat carried into or out of the furnace by gases is a large item in the heat balance. Not only do such problems present themselves in the determination of furnace efficiency, but also in the study of the application of heat is accessory apparatus, such as stoves, regenerators, waste-heat boilers, driers,.etc. The thermal effect of the use of excess air in the combustion of fuel, the theoretical temperatures of combustion, the quantity of heat in hot blast at various temperatures, the effect of hot blast on furnace temperatures, are a few more examples of frequently occurring calorific problems involving gases. So many are the applications of the data on the heat capacity of gases that the subject merits careful study. The heat in a gas maybe due to its heat of combustion, if it is a combustible gas, and to its temperature. The latter is called its sensible heat and is the heat absorbed or evolved by a gas as its temperature is raised or lowered. The heat of combustion of gases is well established and is commonly known. The values can be found in almost any mcdern treatise on metallurgy. The sensible heat of gases has not been so well established because it is in most cases a function of the temperature, and the values of specific heats of gases over a wide range of temperatures have only recently been determined. It is the purpose of this paper to deal with the sensible heat of gases.

Jan 8, 1914

By Walter Murphy

For a number of years the Bureau of Mines has been studying the properties of crude petroleums produced in the United States and the Western Hemisphere. A number of reports have been published giving the analyses of crudes from the different field and regions as made by the Bureau of Mines Hempel method of analysis.4/ Several of these reports include analyses off crude oils from the Rocky Mountain region.5/ The search for new fields and new producing horizons in old fields in the Rocky Mountain region has been quite intensive during the last few years, and a number of discoveries have been made.

Jan 1, 1937

By Todd Smith, Brigitte Lacouture, Greg Anderson, David Lin

"The lead and zinc flotation circuits at Red Dog Mine are preceded by a preflotation circuit which recovers naturally-floating organic carbon contained in the ore. Organic carbon is a potential contaminant in the lead concentrate and is detrimental to lead-zinc selectivity. Prior to 2007, the Red Dog preflotation circuit consisted of a single stage bank of tank cells which floated the organic carbon. The preflotation concentrate was discarded directly to tailings, however, this stream contained significant amounts of lead and zinc. Zinc loss to preflotation ranged from 2% to 6%, relative to the zinc in the mill feed and was proportional to the amount of organic carbon in the ore. The main mechanism of the zinc loss was entrainment.Pilot plant and modeling testwork were conducted to examine the use of a Jameson Cell to clean the preflotation concentrate. Results indicated that up to 90% of the zinc deporting to preflotation concentrate could be returned to the flotation circuit for recovery. The installation of a 5.4 m Ø Jameson Cell with 18 downcomers was initiated in June 2006. The new preflotation cleaning circuit was commissioned in March 2007. Conservatively, preflotation cleaning has resulted in zinc and lead absolute recovery gains of 1.0% and 1.5%, respectively. The $9.1M project has a payback period of approximately 1 year at current metal prices.INTRODUCTIONRed Dog Preflotation CircuitRed Dog Mine is located above the Arctic Circle in northwest Alaska. It is the world’s largest zinc concentrate producer with a production rate of more than 1.0M t/yr. Red Dog is operated by Teck Cominco Alaska Incorporated under an operating agreement with the Northwest Alaska Native Association (NANA). Due to the remote location of the mine, it is accessible only by air, or seasonally, by ocean-going barges.The Red Dog deposit is a rich sedimentary exhalative (sedex) zinc-lead-silver deposit. The major sulphides, in decreasing order of abundance, are sphalerite, pyrite, galena and marcasite. Due to the complex nature of the ore body, run-of-mine ore is blended in stockpiles before being crushed and processed through the concentrator. The average mill feed grade for 2006 was 6.1% lead, 20.6% zinc, 8.0% iron and 0.67 % organic carbon."

Jan 1, 2008

By P. -Q. Wu

A modified electrochemical noise (EN) technique has been applied to monitor corrosion wear of TiN-coated ASP23 steel and TiN-coated AISI 316 stainless steel sliding against corundum in 0.5 M H2SO4. Experimental results show that the EN technique can sensitively detect potential and current variations during a corrosion-wear process. Corrosion-wear mechanisms of TiN coatings depend on their substrate properties. If the substrate is passive, such as AISI 316 stainless steel, the potential and current variations reflect the properties of the coating. Depassivation and repassivation alternately take place on the tribo-activated wear area during the steady-state phase. If the substrate is active, such as ASP23 tool steel, the electrochemical noise is dominated by dissolution of the substrate at the pinholes and the hydrogen evolution all over the exposed surface.

Jan 1, 2005

By W. J. Nokleberg, B. A. Bouley, T. K. Bundtzen

Central Alaska contains a wide variety of pre-, syn-, and post -accretionary mineral deposits that are related to the growth and evolution of the north-west North American continental margin. Withinpre-accretionary terranes, the Devonian-Carboniferous, carbonate platform-hosted MVT zinc-lead deposits of the Kandik and Reef Ridge districts in central Alaska, and the Devonian-Mississippian volcanogenic massive sulfide (VMS) polymetallic deposits of the Bonnifield, and Delta mineral belts of eastern Interior Alaska have received the most exploration interest to date. Syn-accretionary rift related, mafic and ultramafic sill swarms and zoned plutons of the Kluane-Nikolai igneous belt are currently being explored for Ni-Cu-Co-PGE deposits. Syn-accretionary granitoid complexes in the PurcellMountains, Illinois Creek, and Sitylemenkat areas have been explored for uranium-REE-gold, and tin-silver mineralization respectivelyPost-accretionary granitoid intrusions and associated volcanic fields that range in age from mid-Cretaceous to early Tertiary contain significant gold-silver-polymetallic resources in the East Central Alaska, Hogatza, and Kuskokwim Mineral belts. Collectively the three main post-accretionary plutonic/volcanic belts of central Alaska are part of the Tintina Gold Province, where exploration since 1990has identified about 848 tonnes gold.

Jan 1, 2000

By Adrian Moore, Bryan Goodwin, David Fales, David Lam

This paper describes the development of blasting protocols used to successfully create a 450m long blast fractured permeable trench in a basalt aquifer up to 20m thick, at a petrochemical facility in Australia, to enable the remediation of contaminated groundwater.

Jan 1, 2011

By Reginald S. Dean

THE commercial availability of electrolytic manganese has greatly changed the position of manganese as a nonferrous alloying metal. Manganese metal commercially available up to about ten years ago was reduced with either silicon or aluminum and contained substantial amounts of the oxides of these metals and other impurities. Manganese was accordingly not available as an alloy base and its addition to copper-base and nickel-base alloys was generally limited to a manganese content of a few per cent. Several special alloys like manganin contained up to 12 pct of manganese. A method for the production of electrolytic manganese was developed by the Bureau of Mines in 1936 and commercial production was begun in 1938. At the present time (June 1951), production is about 12 tons per day but appears likely to be substantially increased. Electrolytic manganese is deposited from an aqueous electrolyte containing manganese and ammonium sulphates and is marketed as cathode fragments about 1/8 in. thick. The impurities in its composition are as follows: iron, 0.0015 pct; copper, 0.001; arsenic, 0.0005; cobalt, 0.0025; lead, 0.0025; molybdenum, 0.001; sulfide sulfur, 0.017; sulfate sulfur, 0.014; insoluble, 0.002; carbon, 0.002; hydrogen, 0.015. These figures are from Bulletin 463, U. S. Bureau of Mines. From an alloying standpoint, the most important impurity is hydrogen. This can be removed by heating; 95 pct of the hydrogen can be removed in this way without introducing significant amounts of oxygen or nitrogen. Much of the commercial electrolytic manganese used in alloys is dehydrogenited by the producer, since the amount

Jan 1, 1953

By C. J. Konya

When blasts are designed, the assumption is made that all blastholes release near ideal energy. In many types of mining operations, some blastholes release very little useful energy. These blastholes will be called malfunctioning blastholes. Recent studies indicate that on many types of blasting applications, 30 percent of blastholes are malfunctioning. This malfunctioning routinely occurs on the operation and is not an unusual occurrence. This paper will discuss the cause of malfunctioning blastholes and methods to minimize the detrimental effect on both safety and costs.

Jan 1, 1988

By D. Turnbull, J. C. Fisher

ACCORDING to the "reaction-path" modell,2 of martensite nucleation, the shear angle of the embryonic martensite plate must be treated as a variable, and included in any calculation of nucleus critical size. Also, as can be deduced from this model, the interfacial free energy between austenite and martensite does not reach its final value until the shear is completed. It is zero for zero shear angle. However, in order to account for the kinetics of the martensite transformation, some sort of interfacial energy barrier appears to be necessary even with the reaction-path model, for otherwise the volume and the energy of formation of the critical size nucleus both collapse to zero.3 Cohen independently suggested that surface energy could be incorporated into the reaction-path model, with the overall free energy of a martensite embryo being a function of its volume and shear angle.' It is possible to estimate the energy associated with the formation of a critical-size martensite nucleus starting with the reaction-path model and including a surface free-energy barrier. As the dependence of interfacial free energy upon shear angle is unknown, a simple type of dependence will be assumed, with the belief that the true dependence would not lead to appreciably different results. Consider the work required to form a lenticular martensite plate with radius r, thickness t, and shear angle 8. There are three contributions; one being the interfacial free energy, one being the free energy change in the martensite plate, and one being the free energy increase in the surrounding austenite. The interfacial free energy u is assumed to depend upon the shear angle 0 according to the relationship s=s0(?/?0)n [1] where 8, is the equilibrium shear angle and n is an exponent that may lie in the range 0 n 2. The work required to form the interfaces of a martensite plate then is W. = 2pr² s0(?/?0)n [2] The free energy change per unit volume of martensite is composed of two parts, one the ordinary volume free energy ?f1. which is negative, and the other the elastic strain energy G?m²/2, where G is the shear modulus and 7, the shear strain relative to the martensite structure. This expression for the strain energy is valid only when the shear strain ym, is sufficiently small that the martensite is within its linear elastic range. There is no doubt that ym, lies beyond the linear elastic range for embryos that are considerably subcritical. However, for critical nuclei it will be shown that ym, is 1.5 pct or less, within the linear elastic range of martensite. For embryos of nearly critical size, then, the strain energy of the martensite is correctly given by G?m²/2. The shear strain in the martensite is ym, = 8, — 8, and the work required to form the strained martensite is Wm --= (pr²t/2) [?fv + G(?O - ?)²/2] [3] The free energy change in the austenite is entirely that due to elastic distortion. The elastic strain is not uniformly distributed in the austenite, being large near the martensite plate and small elsewhere. Approximately, however, the energy corresponds to a uniform shear strain ya= (?t/2)/r [4] throughout the volume 4pr³/3 surrounding the plate. The work required to strain the surrounding austenite then is Wa = (4pr³/3) (G?a²/2) = (G?²/6) prt² [51 For simplicity, the same shear modulus G is assumed for each structure. The total free energy for forming a plate then is W = W3 + Wm + Wa. = 2pr² s0 (?/p?0)n + (pr²t/2) [?fr+G(?0-?)²/2] + (G?²6) prt2 [6] This expression is correct for nuclei and for embryos of nearly critical size, where, as will be shown, the strain energy in the martensite is correctly given by the expression G (? — ?)². Having W as a function of r, t, and 8, as in Eq. 6, there is a saddle-point where W has a stationary value, W subsequently decreasing indefinitely as the nucleus volume increases along the reaction path. The stationary value of W is the energy of the critical nucleus. The critical nucleus has radius, thickness, and shear angle such that ?W/?r - awlat: = ?W/?p? = 0. Performing these differentiations and calculating the critical nucleus energy, W* = [8192p(G?/6)²;s/27 ?fv4] [7] where a= (?/?0)3n+1[l +G(8"-8)'/2af.]' [7a] and where 8 is to be determined from the equation (1 + 3n/4) + G8(6O - (9)/[Af. +G(6>o-6>)72] = 0 [8] For ?f, near —200 cal per mol or —10" ergs per cc, and 8, near 1/6, as for iron-base alloys, Eq. 8 gives ?0 - ? ~ - (4 + 3n) ?f1./4G0O [9] as the difference between the equilibrium shear angle and the actual shear angle for a critical nu-

Jan 1, 1954

By Waldemar Lindgren

CONTENTS. [ ] I.-CHARACTER OF THE DEPOSITS. 1. Principal Features. IN many schemes of classification and description the term contact-deposit has been somewhat loosely applied to all accumulations of useful minerals (other than those of unquestioned sedimentary origin) which are enclosed between two different rocks. As thus used, the term may include deposits of widely differing origin, and, unless qualified, is not in place in a genetic classification. The present paper deals with a special class of contact-deposits. In many geological provinces, granular igneous rocks, such as granite, diorite and syenite, have broken through and invaded sedimentary rocks. The molten magma may in part have reached the surface and there solidified with relative rapidity as a lava. The largest masses of it, however, did not reach the surface, but cooled very slowly at considerable depth under great pressure, and eventually consolidated into a rock of granitic texture. Uplifts and extensive erosion may have followed; and at the present day, in many places, thousands of feet of material have been removed, bringing to the surface the intrusive granular rocks and their once deep-seated contacts with the sedimentaries which they shattered at the time of in-

Jan 1, 1902

By E. S. Machlin

RECENTLY, there has been much activity in the field of nucleation of martensitic transformations.'- One of the problems in this field is the calculation of the rate of nucleation (number of plates nucleated per unit untransformed volume per unit time) in terms of observable parameters. Two such observable parameters are the fraction of the length of a straight line which corresponds to the portion of the total length of the line that lies in the martensitic phase and the number of plates intersected by this straight line per unit length of line. Because of the necessity of a valid answer to this problem, the following investigation was undertaken. Although this problem has been treated in the past, a derivation of the relation between the volume and linear fractions of martensite will be described to illustrate the method used in the succeeding analyses. It is assumed that the martensitic plates are in the form of a right circular cylinder having a small height relative to the radius of the base. The volume fraction of martensite is then: fr = lXNl,hjAl [1] where f, is the volume fraction of martensite, and Nij is the number of plates per unit total volume having the area Ai and the height hi. The linear fraction of martensite is obtained as follows. For the plates having the particular area A, and height h,, the fraction sin 0 dB* will have their cylinder axes between the angles 0 and 0 + d0 to the straight line along which the linear fraction of mar-

Jan 1, 1953

By P. Scinto, A. Festa

"Orway Mineral Consultants Ltd (OMC) as a consultancy has specialized in comminution circuit design for over 30 years. The OMC Power Model is OMC’s in-house software for specific grinding energy power-based circuit calculations. The model uses ore hardness parameters from testwork, circuit configuration, equipment geometry and operating conditions as inputs. The model combines proprietary and published power models to predict the total specific energy and the specific energy for each of the components of the comminution circuit. This makes the software a valuable tool for the design and validation of JKSimMet modelling or actual plant performance. Proprietary calculations have been developed and reconciled with over 100 plant benchmark projects encompassing a variety of comminution circuit configurations.This paper describes the OMC approach for determining power requirements for the design of grinding circuits. The accuracy of the OMC Power Model software is confirmed through plant survey benchmark performance and comparisons are made with other power modelling techniques. Practical examples are presented, showing applications for single stage SAG (SS-SAG), SAB and SABC circuit configurations.IntroductionConventional low throughput grinding options, consisting of multi-stage crushing followed by rod and ball mills, have almost become obsolete in the move to higher milling rates required to process lower metal grades. Grinding circuits with fully autogenous (AG) or semi-autogenous (SAG) mills, often combined with ball mills are now the standard to meet the high capacity requirement for processing ores with lower metal grades. Simulation software and power-based equation methodologies have been established for designing AG/SAG based grinding circuits and predicting their energy requirements.Presently, power-based equations remain the principal method for designers to size mills and estimate the power requirement for comminution circuits. The technique is simple and it relies on the design to specify a feed and product size plus a hardness index to produce a specific energy value (kWh/t). The desired throughput of the mill is multiplied by the specific energy to give the mill power requirement, which is used in turn to size a mill for dimensions, internal geometry and loading conditions to match the power draw."

Jan 1, 2015

By J. Maiti, A. Bhattacherjee

A literature review on quantitative analysis of mine safely studies revealed that numerous investigators explored a wide- range of techniques, including the investigation of bivariate and multivariate statistical models. Although these researchers addressed various issues of accident causation and prevention, they, in general, did not investigate through the use of quantitative models the multivariate relationships of several factors affecting mine safety in a causal framework. In this paper, systems dynamics and structural equation modeling are presented to simultaneously represent and study the effect of policy variables related to accident prevention along with the factors causing accidents and injuries. The application of these models to a group of mines revealed that current training practices at these mines do not appear to alleviate accident occurrences. It is inferred that the use of these quantitative techniques can provide a new direction of research in mine-safety studies.

Jan 1, 2000

By George A. Lawrence, J. C. McHardy

"Field operating experience with an improved flotation mechanism for handling various flotation applications is covered in this paper. Some aspects of designing flotation machines are also presented. Operating data from extensive testing of commercial size equipment under actual plant operating conditions in the following applications is detailed:1. Copper Flotation2. Fine Coal Flotation3. Base Metal Sulfide FlotationDesign features, design improvements, modifications tested, and overall performance data are reviewed. Some mecl1anical performance data as well as metallurgical performance results are presented. Some comparisons with conventional flotation equipment are reported in this paper.Flotation sizing theories are reviewed and an actual equipment selection ex2.mple is presented. Scale-up criteria used for developing large cells are briefly covered.The paper concludes with a brief review of anticipated future flotation equipment design needs. IntroductionThe mining/milling industry is faced with increasing pressure to cut costs while maintaining product quality and indeed, in some instances, expanding production. For copper, nickel, iron, molybdenum, lead and zinc sulphides, coal, phosphate and some other minerals;, flotation remains the primary method of concentration and recovery of the valuable mineral away from the gangue. Cost reduction in flotation can be broadly focused on two areas:1. Flotation Chemicals2. Flotation Equipment"

Jan 1, 1985

By J. M. Hartman

ONE of the most important subjects to the blast-furnace engineer is a thorough knowledge of the conditions affecting the temperature in the different portions of the furnace. All efforts to decrease the consumption of fuel and improve the working of the furnace must be based upon it, and I may, therefore, be permitted to place before you the results of my observations, and detail the conclusions to which they have led me. Tracing the thermic conditions from below upwards we have at the bottom of a blast furnace making No. 3 iron a temperature of 2900° F., which increases slightly to a point a little below the tuyeres. In the immediate vicinity of the tuyeres the temperature is somewhat lower, owing to the entering blast ; but a short distance above the tuyeres, where all the oxygen of the blast has been converted into carbonic acid, the highest temperature in the furnace must be attained. This carbonic acid is, however, almost as soon as formed, converted by the glowing coal into carbonic oxide, a process which absorbs heat, reduces the temperature, and provides the active agent for the reduction of the iron ores. In its ascent the hot carbonic oxide gradually parts with its heat, first melting the descending iron and earthy materials, which drop down to the hearth of the furnace. The limit of this zone of fusion is rather sharply defined, and the temperature above this zone suddenly decreases. In its further ascent the hot gas drives off the carbonic acid from the limestone, which causes a further absorption of heat, and finally the gas escapes at the top, at a temperature of about 250° F. if the furnace is working well. These changes of temperature 1 have attempted to represent graphically,

Jan 1, 1880