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By B. Tiryaki, M. Ayhan, O. Z. Hekimoglu

Coal shearers are one of the major sources of dust and fines in underground coal mines. Despite considerable research investigations, no comprehensive in situ investigations coupled with extensive laboratory studies on the effects of drum design on generation of dust and fines have been found. It may be for this reason that many major shearer drum manufacturers persistently employ a certain drum-lacing pattern on current coal shearers, which is believed to cause excessive dust and fines in coal mines. This paper shows how respirable dust and fines are minimized under near-identical conditions when the drum lacing pattern was changed from a triple-tracking to a single tracking arrangement. A series of laboratory trials simulating in full scale the practical cutting action of tools was followed by extensive long-term underground trials during production operations in OAL (Cayirhan) Lignite mine in Turkey. The results are presented and discussed, and recommendations for proper drum-lacing in terms of respirable dust and fines are made.

Jan 1, 2002

By S. P. Cashin, D. Kashuba, W. Esplin

Barrick has had 6 autoclaves in operation at the Goldstrike Mine, North of Carlin Nevada, since 1993. The 6 autoclaves operate in parallel, processing slurry of gold ore. The gold in the ore is predominantly associated with, and encapsulated by, sulphide minerals. It is necessary to oxidize at least 90% of the sulphide mineral before the gold can be concentrated using a cyanide/carbon-in-leach (CIL) circuit. Sulphide oxidation chemical reactions occurring in the autoclaves are exothermic; however, the sulphide content in the ore is not high enough to operate the autoclaves in an auto-thermal mode. Heat from steam is added to the first 2 of 5 autoclave compartments to initiate the sulphide oxidation reactions. The steam used to heat the autoclaves is supplied by three boilers, which are fuelled with propane gas. The completion of a few operations and maintenance projects in the last 2 years has resulted in a 50% reduction in propane energy requirements. The autoclave operating temperatures were minimized. Excess oxygen addition was held as low as possible. The material of construction for the compartment walls inside the autoclaves was changed from brick to Titanium metal. And, the speed of the autoclave slurry agitators was increased. All of these activities had an impact on the propane energy requirements of the autoclaves. Presented below is a description of the autoclave systems, and a discussion of how the energy needs were reduced, resulting in significant operating cost savings.

Jan 1, 2004

By M. Laliberte

Gold mine effluents are often characterized by a high concentration of copper, ammonia and cyanide derivatives (mostly cyanates and thiocyanates). While non-acutely toxic to birds, this water can be acutely toxic to fishes and invertebrates. In many jurisdictions, the effluent requires further treatment in order to pass a mandatory requirement to be non-lethal to species like the rainbow trout (Oncorhynchus mykiss) or Daphnia magna. Using actual mine water, it has been demonstrated that aerated moving bed biofilm reactors will biodegrade the ammonia, cyanides, cyanates and thiocyanates, thereby removing the toxicity associated with these components. By removing copper by precipitation the water can be made to pass acute toxicity tests. Tests have demonstrated that the process is stable between 4 and 25 C, with no significant changes in the treated water composition. Once the biomass is acclimatized, the process is resistant to upsets and requires very little inputs except for phosphorus and air. Furthermore, it has been demonstrated that an anoxic reactor can be added downstream of the aerated reactors to remove nitrates if required.

Feb 23, 2014

By M. Laliberte

"Gold-mine effluents are often characterized by high concentrations of copper, ammonia and cyanide derivatives (mostly cyanates and thiocyanates). While not acutely toxic to birds, this water can be acutely toxic to fishes and invertebrates. In many jurisdictions, the effluent requires further treatment in order to pass a mandatory requirement to be nonlethal to species like the rainbow trout (Oncorhynchus mykiss) or Daphnia magna. Using actual mine water, it has been demonstrated that aerated moving bed biofilm reactors will biodegrade the ammonia, cyanides, cyanates and thiocyanates, thereby removing the toxicity associated with these components. By removing copper by precipitation the water can be made to pass acute toxicity tests. Tests have demonstrated that the process is stable between 4 and 25°C, with no significant changes in the treated water composition. Once the biomass is acclimatized, the process is resistant to upsets and requires very little inputs except for phosphorus and air. Furthermore, it has been demonstrated that an anoxic reactor can be added downstream of the aerated reactors to remove nitrates if required.IntroductionThis paper will demonstrate that gold-mine effluents, which are toxic to fishes and invertebrates, can be made nontoxic by using a combination of chemical precipitation and biological oxidation.The typical composition of gold-mine effluents will first be described. The applicable legislation will then be introduced, with its requirement that the treated effluent be nontoxic. The toxicity of a typical untreated effluent will then be confirmed, and the sources of the toxicity identified. After that, a number of toxicity reduction treatment methods will be presented, and the selected method will be described.The lab tests that confirmed the efficiency of the treatment method will be presented, with three phases of testing as the understanding of the process gradually improved. Toxicity test results showing that the water is nontoxic after treatment through the temperature range will be presented."

Jan 1, 2015

By M L. Habner, S D. Munro

In a gold CIP/CIL circuit, adding new circuit water to a final tails thickener and then directing the thickener overflow to the head of the circuit significantly reduces the cyanide concentration at the TSF discharge spigot. The new water is used to dilute the final tails. The dilution and recycle of the tails in this manner may have the added benefit of reducing reagent consumption, in some cases to the point where the reagent savings justify the expenditure and Cyanide Code compliance is achieved at no extra cost. Potential circuit configurations and water recycles can be modelled by splicing a flowsheeting and thermodynamics package, such that changes to the composition of the water can be reduced to chemical speciation with a Gibbs free energy minimisation enabling a complete understanding of the circuit water chemistry.

Jan 1, 2007

In a gold CIP/CIL circuit, adding new circuit water to a final tails thickener and then directing the thickener overflow to the head of the circuit significantly reduces the cyanide concentration at the TSF discharge spigot. The new water is used to dilute the final tails. The dilution and recycle of the tails in this manner may have the added benefit of reducing reagent consumption, in some cases to the point where the reagent savings justify the expenditure and Cyanide Code compliance is achieved at no extra cost. Potential circuit configurations and water recycles can be modelled by splicing a flowsheeting and thermodynamics package, such that changes to the composition of the water can be reduced to chemical speciation with a Gibbs free energy minimisation enabling a complete understanding of the circuit water chemistry.

Jan 1, 2007

By Erik Spiller

"Water and cyanide are highly contentious and conflict issues that many gold mining companies encounter worldwide when dealing with local farming communities, industries and cities competing for water as well as non-government and government organizations. Historically, the gold industry has used processes that required minimum amounts of water to treat and recover gold at locations around the world. Examples are dry grinding for refractory gold treatment in the late 1800s and early 1900s and dry blowing in arid/desert regions to recover alluvial gold. In recent times, new dry processing equipment has been developed and include equipment such as air tables and jigs. Researchers have also modified a Knelson concentrator to concentrate gold from dry solids. High pressure grinding rolls (HPGR) with air classifiers are used to treat industrial minerals and pilot trials have shown that the technology can be applied to comminute refractory gold ores ahead of roasting. Dry stacked tailings is being applied on gold mines to reduce overall water consumption This paper will review the available technologies that in combination may be able to substantially reduce the quantity of water used in the processing of gold ores.INTRODUCTION Water makes up most of the world: the planet is 75% water. Some 97.5% is saline water largely in the oceans, so that only 2.5% is fresh water and useful for human needs. The demand for water is driven primarily by population and concomitant economic growth. Overall, some 70% of the water withdrawn from the environment is used in agriculture, 20% by industry, 7% by households and 3% for mining (Brown, 2003: Maupin et al., 2010). Whilst the water needs of the mining industry are minor compared to those of agriculture, an operating mine requires a relatively large quantity of water on a daily basis to carry out operations. Mining activities are often located in remote, arid environments, with limited access to water. There are a multitude of water resources available to the mining industry for its needs. However, much of this water is saline and over the last two decades there has been a continual decline in water quality (i.e. salinity has increased). In today‘s world, there are enormous internal and external pressures to reduce water consumption at a mine, and/or alternatively use water not considered a resource by the local community. Strategies to meet these objectives can be achieved by implementing good water management practices, by recycling process water, reducing water losses by evaporation and seepage (i.e. by increasing and improving dewatering practices) and employing new technologies such as dry stacking and dry processing (Dunne, 2010)."

Jan 1, 2017

By Graeme Jameson

"In current flotation circuits it is customary to grind the whole of the mill feed to an initial size, which is determined by the liberation characteristics of the ore, and the ability of the flotation machines to float the fully liberated particles. Usually, the product of the primary mill will be ground further in a ball mill, to improve the liberation of the values and hence the product grade, before passing to rougher flotation cells. The final grind size is ore-dependent, and is a balance of factors including the relation between surface liberation and grind size, the grain size of the valuable particles within the host rock, and the feed grade. The P80 for the feed to flotation may be as high as 300 µm for free-milling copper ores, and as low as 53 µm for finely disseminated complex ores. The grinding of the feed ore requires considerable input of energy. The rate of replacement of wear components such as mill liners and media, is proportional to the energy input. The costs of these two factors - energy and wear - are usually comparable and together they form the major part of the running costs of a concentrator, and indeed, of the whole mining operation. The energy required increases as the final grind size decreases. This talk will focus on the reduction of milling costs that could be achieved, using a fluidized bed flotation technology that can give high recoveries at coarse particle sizes. A number of different circuit designs will be presented, using the new technology. Calculations show that reductions in the running costs of a mine and associated concentrator of the order of ten to twenty percent could be achieved."

Jan 1, 2014

By R Higashi, D Maruoka, T Murakami, Y Iwami

The ironmaking industry consumes a large amount of fossil fuel derived carbon as heat source, reducing agent of iron ores and carburising agent of reduced iron. Although the demand for drastic decrease of carbon dioxide emission, carbon is an essential element for smelting process of molten iron. The carbon recycling ironmaking process by circulating CO has been already proposed to achieve carbon neutrality. However, the production of molten hot metal is not considered in this process because sufficient amount of carbon does not dissolve in reduced iron by CO. Therefore, our group has suggested a new carbon recycling ironmaking process which can produce hot metal. In this process, free carbon and iron carbides produced by carbon deposition reaction using metallic iron as a catalyst are used. It is known that only Fe3C is obtained as iron carbide by using CO gas, however, Fe5C2 is also produced by adding H2 gas. The composite agglomerated with these carbonaceous materials and fine iron ore (Deposited Carbon-Iron oxide Composite: DCIC) is reduced and melted in a furnace. It is reported that Fe3C in DCIC accelerates the reduction reaction and melting of the composite. In this study, the effects of iron carbides and free carbon on the melting behaviour of DCIC are investigated. Fe3C, Fe5C2 and free carbon were produced by vapour deposition using porous iron whiskers and CO-CO2-H2 gas. These were agglomerated with hematite reagent at a certain ratio to prepare DCIC samples with and without Fe5C2. The samples were heated up to 1300°C in inert atmosphere. The DCIC containing Fe5C2 completely melted and iron nuggets were obtained after the experiment. This behaviour was not observed in the composite without Fe5C2. This indicates that using Fe5C2 is more preferable than Fe3C for molten iron production using DCIC.

Jun 19, 2024

By Y. S. Sun, P. Gao, Y. F. Li, J. W. Yu

"Oolitic hematite ore is one of the most refractory iron ores in the world. It is not feasible to separate iron minerals from raw ore by conventional mineral processing methods. In recent years, a new technology called coal-based reduction followed by magnetic separation has been employed to produce metallic iron powder from oolitic hematite ore effectively. In order to find the nucleation mechanism, a thermodynamic analysis of the carbothermic reduction of oolitic hematite ore was conducted by the HSC 6.0 thermochemical software. The phase transition and microstructure changes during the coal-based reduction were also investigated by X-ray diffraction (XRD), scanning electron microscopy (SEM) and energy dispersive spectroscopy (EDS). The results showed that the iron oxides in the oolitic ore were reduced to metallic iron following the route of Fe2O3–Fe3O4–FeO–Fe in reduction process. The reduction reactions preferentially occur in the high activity regions to form a higher concentration of Fe elements, and then more supersaturated Fe separated out to form new phase nuclei. Under the action of homo-aggregation and principle of minimum free energy, the iron cores continuously absorbed migratory Fe around and grew into quasi-spherical iron particles embedded in slag phase. INTRODUCTION With the rapid increase of steel industry, premium grade iron ore in China has been gradually exhausted, and its supply has become bottleneck for the development of iron and steel industry (Guo et al., 2015). Therefore, it is urgent to exploit the refractory iron ores, such as oolitic hematite ore. Oolitic hematite ore is widely distributes in France, Germany, America, Canada, Pakistan, China, and among other countries (Maynard and Van Houten, 2004; Li et al., 2011). In China, the explored reserve of oolitic hematite ore is 3.72 billion tons, which accounts for approximately 1/9 of total iron ore resources of China (Gao et al, 2016). Oolitic hematite ore possesses complex composition,"

Jan 1, 2018

By Chengyi Ding, Xuewei Lv, Senwei Xuan, Jie Qiu, Yun Chen, Kai Tang

The influence of SiO2 on the reducibility of the CaO–Fe2O3–SiO2 system was fully examined in this study. The isothermal reduction kinetics of CF and CF8S (CaO/Fe2O3 = 1:1, wt% (SiO2) = 8) were investigated through thermogravimetric analysis at 1123, 1173, and 1223 K with 30% CO and 70% N2 gas mixtures. The reduction of the samples with 8% SiO2 was not only highly accelerated but also proceeded easily. Rate analysis revealed that CF and CF8S reduction occurs in two stages, the Fe3O4-to-FeO stage overlaps with the previous Fe2O3-to-Fe3O4 stage and tends to approach the following FeO-to-Fe stage with the addition of SiO2. The apparent activation energy values of CF and CF8S reduction are 46.89 and 8.71 kJ mol−1. Sharp analysis indicated that CF and CF8S reduction was expressed by the Avrami–Erofeev equation presenting a 2D shrinking layer reaction.

Mar 1, 2017

By Mansoor Barati, Yindong Yang, Yonggang Wei, Bo Li, Alexander McLean

Reduction roasting followed by magnetic separation is an important method for treating laterite nickel ores. In this work, garnierite type laterite ore from Yunnan province in China was reduced by methane under different conditions followed by magnetic separation to obtain the ferronickel concentrate. The phase and microstructure transformations of garnierite in the reduction process were investigated using XRD and SEM-EDS. The experimental results showed that reduction temperature from 873 to 1173 K had only a small effect on content and recovery of nickel and iron. The content and recovery of nickel and iron in the concentrates increased slightly with prolonging reduction time. In addition, the nickel and iron content decreased with increasing methane concentration, but the recovery of nickel and iron increases. A decrease in the iron content and recovery were observed when the dosage of sodium sulfate varied from 5 to 20% at a temperature of 1073 K and reducing time of 90 min, while the nickel content and recovery gradually increased.

Mar 1, 2018

By Chengyi Ding, Xuewei Lv, Senwei Xuan, Chenguang Bai, Kai Tang, Gang Li

Sinter is an important iron-bearing material charged into blast furnace. Calcium ferrite (CF) is one of the most significant liquid phases in fluxed sinter. The isothermal reduction kinetics of CF and CF8M (n(CaO):n(Fe2O3) = 1, wt%(MgO) = 8) with 30% CO and 70% N2 gas mixtures at 1123, 1173, and 1223 K were investigated through thermogravimetric analysis in this study. Rate analysis illuminates that the reduction processes of CF and CF8M are mainly expressed as a typical two-stage reaction. Sharp analysis indicates that CF and CF8M reduction is expressed by the Avrami–Erofeev equation presenting a 2D shrinking layer reaction. The apparent activation energy values of CF and CF8M reduction are 46.9 and 31.8 kJ mol−1 individually based on model-free method.

Mar 1, 2018

By Wing Ka Ng

Substitution of pulverized coal injection by solid biocarbon fuel has the potential to achieve substantial reduction in GHG emissions associated with blast furnace ironmaking. A systematic evaluation was conducted on the performance of solid bio-carbons produced from a single raw biomass source using different pyrolysis technologies. The potential of solid biocarbon in reducing GHG emission of the blast furnace ironmaking process is strongly influenced by its O/C mass ratio. In order to achieve substantial GHG reduction, it is necessary to replace coal injection by solid biocarbon having a low O/C mass ratio. The pyrolysis technology and conditions employed are key parameters influencing O/C mass ratio of the resultant solid biocarbon. Close collaboration between steelmakers and solid biocarbon producers is critical for producing suitable solid biocarbon fuel to replace pulverized coal injection. A successful partnership should pave the way in achieving substantial reduction in GHG emissions of the blast furnace ironmaking process.

By E. Asselin, G. K. Yue

"In the present study, electrochemical measurements, especially cathodic potentiodynamic polarization (CPC) and cyclic potentiodynamic polarization (CPP), have been employed to study the kinetics of ferric reduction on chalcopyrite in aqueous Fe(II)Fe( III)H2SO4 solution at 90 °C and 110 °C. Results from the calculation of speciation in the above solution reveal that, at both 90 °C and 110 °C, the real Fe3+/Fe2+ ratio is much lower than the nominal ratio, which leads to smaller VALUES (for the reversible potential of the ferric/ferrous couple. Exchange current densities of the ferric/ferrous couple were obtained and are on the order of 106 10 5 A/cm2 a t 9 0 °C and 1 07 10 6 A/cm2 at 110 °C, respectively. The Tafel slopes from the CPP method are higher than those from the CPC method at both 90 °C and 110 °C. The transfer coefficients were about 0.360.59 at 90 °C and 0.370.65 at 110 °C. Based on the exchange current densities obtained, the rate constant kc was calculated. It was found that kc at 90 °C was slightly larger than at 110 °C, especially for the results obtained by the CPC method. The average rate constants are 862.17 cm/s for CPC and 65.30 cm/s for CPP at 90 °C, and 477.69 cm/s for CPC and 63.53 cm/s for CPP at 110 °C for nominal Fe3+/Fe2+ ratios ranging from 1:1 to 1000:1. INTRODUCTIONChalcopyrite is one of the most abundant copper-bearing minerals, accounting for approximately 70% of the world’s known copper reserves (Wang, 2005). Most of the currently proposed leaching processes are based on sulfuric acid with ferric ions and/or dissolved oxygen as oxidants. Some of these processes have been piloted in attempts to implement these technologies on an industrial scale, such as Activox (U.S. 5,232,491, 1993), Nenatech (U.S. 5,993,635, 1999), Dynatec (U.S. 5,730,776, 1998), AAC/UBC (U.S. 6,503,293, 2003), Total Pressure Oxidation (U.S. 5,698,170, 1997), CESL (U.S. 5,874,055, 1999) and Galvanox (U.S. 7,846,233, 2010).To date, much effort has been devoted to understanding the kinetics and mechanisms of the oxidative processes (Biegler & Horne, 1985; Córdoba, Muñoz, Blázquez, González, & Ballester, 2008, 2009; Ghahremaninezhad, Asselin, & Dixon, 2010; Hirato, Majima, & Awakura, 1987; Hiroyoshi, Kuroiwa, Miki, Tsunekawa, & Hirajima, 2004; Warren, Wadsworth, & El-raghy, 1982) or the cathode reduction reactions of chalcopyrite itself (Arce & González, 2002; Biegler & Swift, 1976; Biegler, 1977; Nava, González, Leinen, & Ramos-Barrado, 2008), by chemical or electrochemical methods. However, as the most important cathodic processes in chalcopyrite leaching systems, the cathodic reductions of ferric ions or dissolved oxygen are worth measuring due to their effects on the overall leaching rate, as predicted by mixed potential theory."

Jan 1, 2012

By Margaret E. Harris

The formation of nitrogen oxides, the decay of carbon monoxide, and the concentrations of residual hydrocarbons in the secondary combustion zones of propane-air and methane-air flames were studied in three enclosed burners designed to simulate gas appliances such as space and water heaters. The effect of thermally or chemically perturbing the combustion gases above these flames was determined theoretically and by means of factorial experiments. The concentrations of air pollutants and the flame temperatures were measured, and the effects of ingress of secondary air into the primary and secondary combustion zones, flame stability, burner port surface, and simulated recycling of combustion gases were noted. Based on this study, four principles were derived that, applied to gas appliance design, will insure minimal emissions of these air pollutants: 1. Oxides of nitrogen can best be kept low by depressing peak temperatures to about 3,050° F. 2. If this is not possible, then the secondary combustion zone should be cooled rapidly to this temperature. 3. Carbon monoxide concentrations can be limited by rapid induction of secondary air. 4. Hydrocarbon concentrations can be controlled by designing for well-seated, stable flames. In addition, empirical reaction rates were derived that predict the concentrations of nitrogen oxides and carbon monoxide in the primary and secondary combustion zones.

Jan 1, 1970

By Jörn E. Fischer

Cracks were observed at the top of spray formed billets for some alloys. In this paper the effect of process parameters on crack probability is discussed. After the spray run the thermal conditions at the top of the billet change rapidly. A hot spot with melt residue] can occur inside the billet and shrinkage will be suppressed due to the solidified surface of the billet. In this situation hot cracks can be initialized. The temperature history of the billet during and after the spray run is calculated. The temperature difference between core and outside diameter of the billet is taken as indicator for hot crack probability. The effect of different process parameters on the hot crack probability is discussed. The results lead to a better understanding of the process. The calculations are taken as a tool to find process parameters in order to reduce or avoid hot cracks in spray formed billets.

Jan 1, 1999

By G. G. Krishna Murthy

Interactions of wustite and hematite pellets with Fe-C melts were investigated experimentally by examining the effect of pellet size and bath temperature on the reduction rates of the pellets. Experiments were carried out in the temperature range 1473-1923 K. A flash X-ray equipment was used to obtain images of the melt surface during its reaction with an oxide pellet. The reduction rates were calculated using experimentally estimated time for complete reaction of a pellet, tR. The value of tR was determined employing gas mass flow rate, or pressure rise in the sealed furnace, versus time traces. The value of tR was the same for both wustite and hematite pellets of equal weight at or above 1723 K and it decreased with increasing melt temperature. But the reduction rate was substantially lower at temperatures below 1651 K, the melting point of wustite. The rates were defined as the rate of moles of oxygen removed from the pellet per unit area of bath free surface and it was found that the bath surface area affects the reduction rates. It was established that the overall reaction followed nth-order rate kinetics with the values of n falling between 0.3 and 0.4 for temperatures above the melting point of wustite. The apparent activation energy determined using the apparent rate constants obtained from the experimental data was 76.2 + 27.7 kJ/mol-O.

Jan 1, 1992

By Martin Froehling, Curtis Mohns, Rahim Rajwani, Michael Fullam

"The Kemess mining and milling complex is operated by Northgate Minerals Corporation and is located in the mountains of north-central British Columbia, 430 kilometers northwest of Prince George. The complex consists of the Kemess South open pit mine and a 52,000 tonnes per day mill. Hypogene, supergene and leach cap ores are processed using conventional crushing, grinding, and flotation techniques to produce gold-copper concentrates with minor by-product silver value.The rougher flotation concentrate is re-ground in closed circuit before being cleaned to final concentrate grade. Historically, gold losses in the cleaner tailings were primarily associated with pyrite, with only minor amounts of free gold. More recently, with the processing of high pyrite ores, the major losses were found to be free gold less than 10 microns in size. The fine grind required for the high pyrite ores resulted in over-grinding of the free gold particles in the regrind circuit, contributing to poor flotation characteristics. It was thought that if this free gold could be recovered by gravity at coarser size in the regrind circulating load, the free gold losses in the cleaner tailings could be reduced.Preliminary test work carried out by Knelson Gravity Solutions indicated the potential for free gold recovery in the regrind circuit. A Knelson XD-30 Concentrator was installed with variable speed capability, and a series of optimization tests were carried out. In addition, cleaner tailings were characterized with gravity operating and with gravity not operating to determine the overall gold recovery benefit. The single Knelson XD-30 installation recovers ~10% of the total gold in the regrind circuit and increases overall recovery by 3% when high pyrite ores are processed."

Jan 1, 2007

By Arthur B. Foote

THE ball mill has superseded stamps for the reduction of gold ores in most of the recently designed plants, partly because stamps are not suited to die fine grinding required for flotation, and partly to save space and construction cost. Experience with these new mills seems to confirm the idea that it is more difficult to recover bold from a ball-mill product than it is when stamps are used for crushing. Microscopical examination of particles of gold that have passed through a ball mill show that the gold has been pounded into exceedingly thin flakes coated with a talcy surface, making it more difficult to save by amalgamation, flotation, or concentrating. Bearing in mind that gold is extremely malleable, it seems reasonable that this should be so. Contrast the action in a stamp battery, where the crushing is by impact in the presence of mercury, to the grinding and rolling under pressure in a ball mill. It is well known that a better recovery can be obtained by panning or amalgamation from a sample of ore that has been crushed by mortar and pestle than from one reduced on a bucking board.

Jan 1, 1937